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By C. M. Haight

The mine blacksmith and drill-steel sharpening shop at the Franklin mine of the New Jersey Zinc Co. is on the surface, adjoining the main shaft. It is a brick building, 51 by 30 ft. inside dimensions, with a corrugated iron roof and concrete floor. .About two-thirds of the steels used are brought up through the main shaft; the remainder through two auxiliary shafts and are carried to the shop by team or truck. At each shaft sufficiently large stocks of sharp drills of various typcs are kept to replace all the dull steels that may be sent up; while at the shop enough sharp drills are kept to replace the dull steels in the stocks kept at the auxiliary shafts. Though this entails keeping a larger number of drill steels in service than might be required by other methods, there is less hurry in the sharpening shop and the steels have a longer period of rest. If a rest from use enables steel to recover from fatigue strains, this method secures that rest. The forms shown in Fig. 1 are supplied to each shift boss in books of 100. These are used to requisition sharp drills from the shaft stocks and cnable the men at the tops of the shafts to sort out the required sizes and to hare them ready to be sent to the various shaft stations. At one of the auxiliary shafts, the dull drills are placed on a platform from which they are dumped into the body of the wagon that takes them to the shop.

Jan 1, 1927

By AIME AIME

The following report is published for the information of the members Meetings. Two meetings for the reading and discussion of papers, etc., have been held during the year 1906-namely, the Ninetieth meeting, held February 21st to 24th at Bethlehem, Pa., and the Ninety-first meeting, held July 24th to 27th in London, England, in connection with the meeting of the Iron and Steel Institute. The proceedings of these meetings, including descriptions of the entertainments and excursions connected therewith, have already been published and distributed to the members of the Institute; the Bethlehem meeting in Bi-Monthly Bulletin, No. 9, May, 1906, pp. 497 to 508, and the London meeting in Bi-Monthly Bulletin, No. 12, November, 1906, pp. 809 to 908. The proceedings of the London meeting were published separately and distributed among the many individuals and concerns whose cordial friendship and delightful hospitality contributed so largely to the success of the visits in London and environs and the excursions in England, Scotland and Germany. At the Bethlehem meeting, 37 papers and discussions were presented and the names of 198 members and guests were regis¬tered at the Institute headquarters ; this number, however, does not represent all who were present at the sessions and excur¬sions. At the London meeting, 26 papers and discussions were presented, not including 5 of the papers and discussions read before the Iron and Steel Institute at the joint session with the American Institute of Mining Engineers, which, by a mutual agreement between the Councils of the two Institutes, have been selected to be republished in the Bi-Monthly Bulletin and Transactions of the Institute. The number of members and guests of the two Institutes registered at the London headquarters was 908, and of this number 131 were from America. This list does not include the names of all who

Mar 1, 1907

By E. J. KENNEDY

METHODICAL AND EFFECTIVE: thus may be characterized the fall meeting of the Iron and Steel and Institute of Metals Divisions at the Hotel Statler, Buffalo, N. Y., on Oct. 4 and 5. Approximately 200 registered. Interest centered 1111 the technical papers which were keenly followed, with substantial discussion: During the entire week of the National Metal Congress (Oct. 3 to 'i), there were also other sessions of interest' to metallurgists, directed by the America11 Society for Steel Treating, American Welding Society, Production Division of the Society of Automotive Engineers, the Iron arid Steel and Machine Shop Practice Divisions of the American Society of Mechanical En,' the Wire Association and the American Drop Forging Institute. The National Metals Exposition at the 174th Regiment Armory was well attended.

Jan 1, 1932

By G. Van Esbroeck, M. Van Weyenbergh

The copper mines of the Katanga region in the Belgian Congo lie along the same mineralized belt as those of Northern Rhodesia. There are two distinct types of deposits in that belt, the dolomitic and the quartzitic. In the dolomitic type the metallic minerals are found in a dolomitic limestone country rock. Through oxidation and enrichment, large deposits of high-grade oxidized copper ores were formed, some of which are being worked by the Union Minière du Haut Katanga. Such deposits are all in Katanga. The quartzitic deposits are found in Katanga and Northern Rhodesia. Here the primary mineralization occurred in sandstones and quartzites, in horizons stratigraphically lower than the dolomites. These rocks are not favorable to secondary enrichment and the oxidized zone is poorer than the sulphide zone, which alone contains the ore. There are also in Katanga a few ore-bearing veins. One of them is being worked in the Prince Leopold mine at Kipushi. Prince Leopold Mine The Kipushi ore body fills a fault vein and invades the wall rocks more or less extensively by metasomatic replacement. The vein strikes north-northeast and south-southwest and dips to the west. The variable dip averages 60' to a depth of 300 ft., and more uniformly 80' below 400 ft. In the upper section the average width is 200 ft. and below the width is about 150 ft. The length of the ore body is 2300 feet. The eastern side of the vein—the footwall—is dolomitic and contains irregular masses of ore sometimes extending 100 ft. or more from the main

Jan 1, 1940

By M. Van Weyenbergh, G. Van Esbroeck

The copper mines of the Katanga region in the Belgian Congo lie along the same mineralized belt as those of Northern Rhodesia. There are two distinct types of deposits in that belt, the dolomitic and the quartzitic. In the dolomitic type the metallic minerals are found in a dolomitic limestone country rock. Through oxidation and enrichment, large deposits of high-grade oxidized copper ores were formed, some of which are being worked by the Union Minière du Haut Katanga. Such deposits are all in Katanga. The quartzitic deposits are found in Katanga and Northern Rhodesia. Here the primary mineralization occurred in sandstones and quartzites, in horizons stratigraphically lower than the dolomites. These rocks are not favorable to secondary enrichment and the oxidized zone is poorer than the sulphide zone, which alone contains the ore. There are also in Katanga a few ore-bearing veins. One of them is being worked in the Prince Leopold mine at Kipushi. Prince Leopold Mine The Kipushi ore body fills a fault vein and invades the wall rocks more or less extensively by metasomatic replacement. The vein strikes north-northeast and south-southwest and dips to the west. The variable dip averages 60' to a depth of 300 ft., and more uniformly 80' below 400 ft. In the upper section the average width is 200 ft. and below the width is about 150 ft. The length of the ore body is 2300 feet. The eastern side of the vein—the footwall—is dolomitic and contains irregular masses of ore sometimes extending 100 ft. or more from the main

Jan 1, 1940

By Donald H. McLaughlin

ZEST in the search for new supplies of metallic ores and petroleum is difficult to maintain with stocks of raw materials accumulating and with over- production rightly or wrongly blamed for most of our industrial troubles. Enthusiasm over new methods of prospecting has consequently diminished and exploration programs in many fields have been drastically cur- tailed. Very naturally the effects have been felt with particular keeness in highly specialized phases of the work, such as geophysical prospecting, for the feverish scramble for new areas in which to drill beneath magnetic or electrical anomalies has abated, and there is even a lull in the firing in the battle over salt domes on the Gulf Coast. It is unquestionably an unhappy time for many of the organizations dependent on this work, but in certain respects the situation may be beneficial in correcting unsound practices that .have grown up through the necessity for haste and for quick decisions. With the restriction of the field to the more competent workers by the harsh rule of survival of the fittest, a general improvement, not only in technique but in business methods, can be expected which may go far in restoring public confidence. Geophysical prospecting in the short period that it has been utilized in North America, has suffered severely both from over-popularity and from excessive abuse. The former evil now

Jan 1, 1931

By Elmars Ence, Harold Margolin

A. J. Goldak and J. Gordon Parr (University of Alberta) —While we appreciate the difficulties involved in any investigation of this system, and we wish to congratulate the authors on their comprehensive efforts, there are several aspects of their paper that are not clear to us. 1) If the 6 phase is, in fact, isomorphous with Ti3Sn, it should not be designated Ti,Al. An 4B structure cannot be isomorphous with an ordered structure of the DO type. Anderko et al.' suggested that the structure be called Ti3A1, and we believe this to be a more reasonable designation. The confusion is compounded when the authors designate the phase Ti3Al—yet the composition limits of this phase (according to their diagram) are approximately Ti,A1 and Ti,Al. Further, while Pietro-kowsky is said to have pointed out that the structure of the 6 phase is isomorphous with Ti3Sn, we understand that this was only a suggestion that was not confirmed. Margolin and Ence do not appear to have presented any evidence to support the suggested isomorphism. 2) In referring to the precipitation observed microscopically during the initial breakdown of high A1 a during quenching, the authors state that "such precipitation is not likely to be detected by Debye-Scherrer techniques". We suspect that if the precipitate was visible by optical microscopy, it must have at least one dimension of the order of 1 or greater, and would therefore be readily detected by diffraction techniques. 3) The authors state: "The basis for this suggestion (that 6 forms peritectically P + L—6) is the observation that the ,'3 + 6 and 6 phase fields were

Jan 1, 1962

By R. J. Wysor

Since the outbreak of the European war, few problems of raw-material supply have commanded more nation-wide attention than potash. It is well known that before the war the domestic production of potash was an insignificant percentage of the imports. The average annual importation of raw potash salts for several years prior to 1914 was slightly over 300,000 net tons, and of other potash manure salts about 700,000 tons. The sudden and almost total cessation of these imports created a unique and stringent situation. Methods of recovering potash from feldspar and other native mineral sources, and reclamation from hitherto waste products, have received a marked impetus. Soaring prices have been a keen incentive to research and industrial development. Although several brief articles concerning the possibilities of potash recovery have appeared in recent trade journals, it still may be a matter of some surprise to the average technical mind that potash might be reclaimed as a profitable byproduct in the manufacture of pig iron. At the present time it may be of special interest to present some data, largely of a technical nature, on this subject. A fairly thorough search of the literature reveals a number of articles concerning salts of the alkali metals in the blast furnace. Many of them deal with the theoretical rôle of alkali cyanides in the furnace. Several discuss the probable effect of the alkalies on the furnace brickwork. Two or three ambitious patents have been granted for reclaiming potash or other products from blast-furnace gas. However, we have heard nothing further as to the practical application of these patents. Little literature of importance has appeared during the last 10 years concerning alkalies in blast-furnace practice, and I have discovered no record whatsoever as to the actual sale or commercial disposal of flue dust for its potash content until this was inaugurated in our plant at Bethlehem. A bibliography of the most important articles discovered is appended. About 4 years ago, in the course of investigating blast-furnace stove efficiencies, I analyzed the fine, yellowish fume of which a considerable

Jan 1, 1917

By Wolfgang Hermann

The ancit-process of the Eschweiler Bergwerks-Verein, an affiliate Company of ARBED S.A., Luxembourg, was developed in its principles by the Dutch State Mines, Limburg, in the mid-50's with the aim of producing a smokeless domestic fuel on the basis of low volatile hard coal. In 1967, the EBV erected a 10 t/h large- size pilot plant along these principles in the Aachen mining district and further developed the acquired know-how. In hardly 10 years more than 650 000 t of ancit briquettes with unit weights between 17 and 100 grams were produced and in the meantime stream rates of over 85 % of the yearly utilization hours have been reached. In more than 10 blast furnace trial runs involving substitution rates up to 100 p. c. ancit have proved the suitability of ancit as a metallurgical fuel both for a burden with a high and a poor iron content. Construction work for a minor process-step to enable processing of coal with a higher content of volatile matters is finished and first results are expected by summer this year. A pilot plant with a 30 t/h capacity working on a low-volatile coal basis is being planned. Engineering and construction are supervised by MecanARBED. Luxembourg, as responsible contractors.

Jan 1, 1977

By W. E. Bruce, R. J. Morrell, James Paone

This chapter is a partial summary of an investigation by the U.S. Bureau of Mines (USBM) made in response to a recent request by the Assistant Secretary, Water and Power Development of the U.S. Department of the Interior, for a comprehensive review of horizontal boring technology related to underground power-line installation. This review is concerned primarily with 3 to 36-in. diam horizontal holes from several hundred to several thousand feet in length. An electric power system consists of three essential functions: generation, transmission, and distribution. The generation component includes the generating station and step-up substation; the transmission component includes the transmission line and stepdown substation at typical voltage levels ranging from 69 to 345 kv for underground lines and 69 to 500 kv for overhead transmission lines; the distribution component includes the distribution line carrying the commonly used voltages of 120, 208, and 240 v, the distribution transformer, and the line to the service customer.l The Federal Power Commission reports that 1600 miles of underground transmission lines of 60 to 345-kv capacity are now in service in the United States. By 1980, it is estimated that approximately 3000 miles of trans- mission lines will have to be placed underground.2 The cost of burying electric transmission lines underground in metropolitan areas, approximately 30 miles from the center of each of the most populated cities, will exceed $1.5 billion with the methods now available. The wide spread in unit costs of overhead and underground transmission lines narrows as right-of-way costs for overhead transmission in congested areas increase. Improved methods of underground emplacement will certainly reduce this difference in costs and even further favor underground emplacement. Also, in the future, overhead transmission lines will probably not be permitted in highly built-up areas.

Jan 1, 1970

By E. W. Krampert

The oil industry was very active in Wyoming in 1936, in contrast to the several quiet years preceding. Production for the year again increased about 7 per cent, following an 8 per cent increase in 1935. During the year, 85 oil wells were completed, compared to 55 in 1935. Three gas wells were completed in 1936, and three in 1935. At the end of the year 45 wells were drilling, as compared to 55 drilling at the end of 1935. Two new oil fields and one new gas field were discovered in Wyoming during the year. In the year 1936 Wyoming produced 14,475,431 bbl. of oil, a gain of 1,001,081 bbl. over the 1935 figure of 13,474,350. This increase, in the main, was due to the taking of light oil from the newly developed Sundance sand pool at Lance Creek, and from the new fields, Quealy and Medicine Bow, which contributed their first production this year. Most of the other old fields either were partly shut in in competition with the new fields, or showed a normal decline. This was especially noticeable in the black-oil fields, where production had been slowly but steadily increasing for a number of years. During 1936 black-oil production decreased approximately 100,000 bbl. as compared to 1935. The only black-oil field to show marked increase was Garland, which managed to take the market from less favorably located fields. The production of Wyoming in 1936 was 11,358,757 bbl. of light oil and 3,116,674 bbl. of black oil. Salt Creek, with a production of 6,015,000 bbl., is still the largest single oil-producing field, accounting for 42 per cent of the year's total production. Lance Creek, an old field, with newly developed production in the Sundance sand, reached second place this year for the first time in its history, with a production of 1,878,486 bbl. Lance Creek production will almost equal that of Salt Creek in 1937, if the present rate of production is maintained during the coming year. Oregon Basin, a large black-oil field, held second place in 1935, but owing to a greatly lessened demand for its production, dropped back to third place in 1936. This was in part due to greatly increased light oil available, and in part to the greater supply of black oil available at

Jan 1, 1937

By R. T. Hukki

IT is no great exaggeration to say that present grinding practice and economics are largely determined by lining design. A record of outstanding liner wear can be achieved with any liner surface pattern that will positively lock the outer layer or layers of grinding media. With no slippage, lining wear is bound to be slight. At the same time, popular practice calls for tumbling loads of about 50 pct of mill volume to obtain maximum grinding capacity. Innumerable parallel grinding investigations have verified that optimum speed for such a mill lies within 70 to 85 pct of theoretical critical speed. If the mill is speeded up to 100 pct of critical, little or no grinding can be accomplished. Earlier Work on Supercritical Grinding: First investigations concerning grinding at supercritical speeds seem to be very old. Remarkable work on the subject has been performed by White,&apos; and his experiments have been described and summarized by Richards.&apos; White&apos;s contributions seem to have passed unnoticed by Fahrenwald,2,3 whose extensive experiments have been well presented, yet apparently very little appreciated. Additional work on grinding at supercritical speed has been reported, e.g., by Lewenson and Tscherny,&apos; USSR; Anselm and Grunder,&apos;, Germany: and Rose and Evans." Great Britain. Subcritical and Supercritical Speeds: In the formula of the critical speed given in the textbooks of mineral dressing, no factor indicating the coefficient of friction is generally included. If this factor = 1.0, which is equivalent to grinding conditions in a mill provided with heavily ribbed lining, the formula of the critical speed holds as such and grinding is possible at subcritical speeds only. In a mill equipped with a smooth or relatively smooth lining, the numerical value of the friction factor in the denominator of the formula becomes <1.0, indicating that grinding at supercritical speeds should be possible in such mills. It has been recently shown by the author&apos; that a wide supercritical speed range will become available for grinding—and especially for fine grinding—if the basic conditions within the mill have been selected properly. Mathematical analysis of mill dynamics at supercritical speeds&apos; has indicated that the mill speed may be increased if: 1) total mass of grinding medium decreases, 2) mass of the individual grinding piece increases, and 3) the coefficient of friction between the outer layer of medium and the mill lining decreases. It is obvious that the mass of the individual grinding piece will be affected by its shape and by its specific gravity. The coefficient of friction decreases with: 1) increasing smoothness of liner surface, 2) increasing roundness of the grinding piece, 3) increasing fineness of material to be ground, 4) decreasing pulp density, and 5) decreasing hardness (abrasiveness) of the mineral to be ground. In a mill equipped with heavily ribbed lining, practically no size reduction will take place between the lining and the outer layer of tumbling medium, because slippage is prevented. In the overwhelming majority of today&apos;s mills the grinding accomplished is by virtue of cataracting and/or cascading media with some action within the tumbling charge. Sub-critical speeds only can be applied. In grate-type or peripheral discharge mills equipped with a smooth lining a tumbling load of any kind of common medium occupying about 50 pct of mill volume will behave in such a way that practically no slippage will take place. This has been verified in experiments run by the author in the laboratory and in pilot plant mills&apos; equipped with smooth lining. Again, the operation is limited to the subcritical range. In a mill with a smooth lining, grinding will be possible either at subcritical speeds or within a wide supercritical speed range as soon as the basic requirements for a desired speed are fulfilled. In a mill operated at supercritical speed, any point on the liner surface proceeds at a speed greater than that indicated by the formula of the critical speed, while any grinding piece situated in the outer layer of the medium against the lining proceeds in the same direction at a speed less than that indicated by the critical speed. This speed difference produces a very effective attrition grinding zone between the liner surface and the outer layer of the medium. The share of attrition grinding of the total grinding accomplished increases rapidly with increasing speed in the supercritical speed range. The smooth surface of the mill lining may be well illustrated by the surface of a bucking board. The outer ball layer may be similarly represented by the lower surface of a muller. If the bucking board, the material to be ground, and the muller all proceed in the same direction at the same speed, no grinding will result. This is the general situation in the mills of today. If, however, the muller is pulled with respect to the board, moving or stationary, grinding will be accomplished effectively. Although grinding in today&apos;s mills is primarily the result of cataracting and/or cascading media, the principal place of grinding at high supercritical speeds will be the attrition zone.&apos; In the mineral dressing laboratory of the State Institute for Technical Research. Helsinki, Finland, large quantities of different ores have been ground in a pilot plant ball mill (3x3 ft) at a top speed about 230 pct of the critical. With the same mill, theoretical investigations concerning the grinding characteristics of ball mills equipped with a smooth lining have been carried out up to the speed of 313 pct&apos; of the critical. With a smaller mill, the grinding characteristics with a variety of grinding media have been investigated at speeds up to 2000 pct of the

Jan 1, 1959

E. L. BLOSSOM, New. York, N. Y.-That this paper deals with a real problem is illustrated by a statement made to the speaker a few months ago by the manager of a silver-lead property: "Our ore contains 60 per cent. silica on which the smelters are penalizing us heavily, and they are not able to take all our stuff even then. We have simply got to find some method other than smelting for getting the silver and lead out of our ores. We have had experts down here and have investigated flotation, leaching and volatilization, of which processes the latter is by far the most promising. We are continuing investigation." I was somewhat concerned in certain chloride volatilization experiments which we carried on in Denver in 1903, working on a carload of ore from Butte. Now being connected with the Cottrell process, by which method the fumes will undoubtedly he recovered if they solve the volatilization problem, I have looked into the matter again, and believe it can be developed into a commercial process. On a laboratory scale, working With silver-lead ores, experimenters have obtained a volatilization of over 90 per cent. of the total values in 30 to 60 min, on quantities of 1 to 5 lb., using a roasting dish and a muffle. That looks very good, and would be entirely commercial. On attempting to duplicate these results in a reverberatory furnace, they found that it is more likely to take 8 hr. firing to yield 75 per cent. extraction, which is not commercial at all, as it requires too much time, labor, and fuel, and the whole process is too slow. Of course it can be done, because we have done it in the laboratory.

Jan 1, 1918

By Edwin Ludlow

For many years, coal has been known to exist in the valley of the Sabinas river, in the State of Coahuila, and for about 15 years it has been worked by the Mexican International Railway Co., which opened mines for its own fuel-supply when the railway was built, and has been mining continuously ever since. The coal-basin worked by this company lies along the Sabinas river, principally on the SW. side. It covers a wide area and has been carefully prospected, but only in a few parts have the beds been found sufficiently thick and regular for economical operation. This Sabinas basin was considered the only one in this field, until, in the Spring of 1899, the Las Esperanzas basin was found by Messrs. J. L. Elliot and E. D. Peters, who were in Mexico looking after some copper-properties. After a further examination by Mr. James T. Gardiner, of New York, the Mexican Coal & Coke Co. was organized in June, and active work was begun on Nov. 5, 1899. The coal-field of Las Esperanzas, situated in the State of Coahuila, about 85 miles SW. of Eagle Pass, Texas, adjoins and runs parallel with the Sabinas basin—an anticlinal separating the two outcrops by about 2 miles. Fig. 1 is a sketch-map of a part of the basin. The coal-formation is in the Upper Cretaceous, and corresponds with the Laramie measures of the United States. The following geological section, taken in Nogales canon, shows the measures between the massive limestones forming the Sierra de la Santa Rosa and the outcrop of the coal: Section. Feet. Limestone,...Unknown. Blue shale,.........110 Shale and sandstone,.. 60 Bed of fossiliferous shale,. 60

Jan 1, 1902

By Lawrence K. Jetter, Robert F. Mehl

SOME advances have been made recently in the theory of the kinetics of precipitation from metallic solid solution despite the complexities of the problem, but there is surprisingly little quantitative information available on rates of precipitation with which to test present theory or upon which to build new theory.? Attempts to measure rates of precipitation have been beset with difficulties which originate both in the anomalies that attend the changes in the properties selected for measurement, as, for example, in electrical conductivity or mechanical properties, and also in the complexities in the structural changes in some systems, as in the aluminum-copper system. If the functional relationship between the quantity of precipitate and the value of the selected property is uncertain, measurements of that property are obviously of uncertain utility in determining absolute rates of precipitation. If the precipitating system passes through one or more transition stages, the interpretation of measured rates lacks definiteness; in such cases, it is necessary to measure the rate of reaction in each separate stage, if the rate measurements are to have any significance from the kinetic point of view.* These difficulties cannot be wholly avoided, for at least in some measure they are characteristic of the process of precipitation from solid solution, but they can be minimized both by the selection of a simple system and by the use of a method of measurement relatively free from anomalous behavior. The present paper reports dilatometric measurements on the rate of precipitation of silicon from the aluminum-rich solid solution in the aluminum-silicon system-a study that approaches the desiderata of simplicity. The aluminum-silicon system offers advantages from these points of view: there are no known anomalies in the property changes accompanying precipitation ;2,3,4 there are no known transition stages; the precipitate is an element, not a compound, and both matrix and precipitate are cubic, and while the structures are not identical (the solid solution is face-centered cubic and the precipitate is diamond cubic), the system at least approaches the ideal case of a crystal precipitating from a matrix of the same crystal type-the ideal case which Becker5 assumed in treating the precipitation process theoretically. The dilatometric method, while not free from disturbing effects, as the subsequent dis-

Jan 1, 1942

By J. J. Frawley, W. J. Childs

The dynamic nucleation of supercooled bismuth and Bi-Sn alloys has been studied over a frequency range of 15 to 20,000 cps. For low-frequency vibration, a minimum vibrational energy was required for enhancement of nucleation. Above this critical energy, the dynamic supercooling was less than static supercooling showing that vibration promoted nucleation. The amount of dynamic supercooling continued to decrease with increasing vibrational energy until a minimum or threshold value was reached. This minimum value of supercooling for nucleation remained constant joy all further increases in vibrational energy. For higher frequencies, similar results were observed. This behavior has been related to the necessity of cavitation for dynamic nucleation. When a liquid is cooled to a temperature below its equilibrium melting point, the solid phase is more thermodynamically stable. However, for solidification to occur, a two-step process, nucleation and subsequent growth of the solid phase, must occur. When a liquid is supercooled, that is cooled below the equilibrium melting point, the controlling process for solidification to begin is the rate of nucleation. Once nucleation has occurred, the solidification process is controlled by the rate of growth. Nucleation can be induced by two factors: either by a catalyst or by the use of mechanical shock. Numerous investigators1-4 have studied the effect of nucleation catalysis but much less systematic study has been made of nucleation by mechanical shock waves. The influence of vibrations on grain size in castings and ingots has been studied by many authors but no clear understanding of the mechanism or accurate prediction of the effect has been presented.5 It would be intuitively expected that the further the departure from equilibrium (i.e., the greater the supercooling), the easier it would be to induce nucleation. This has been quantitatively demonstrated both by walker6 and later by Stuhr,7 that the greater the degree of supercooling the easier it is to nucleate by a shock wave. Stuhr also attempted to obtain the mechanical energy required for nucleation of bismuth as a function of supercooling. He vibrated a crucible containing supercooled metal at low frequencies and various amplitudes and noted the corresponding dynamic supercooling obtained. The amount of supercooling was inversely proportional to the mechanical energy applied. Limitation of his experiment was the problem of the confinement of the liquid in the crucible without splashing and minimizing other unwanted modes of vibration. Tiller et al.8,9 did similar work on tin and Sn-Pb alloys using an electromagnetic stirring device. Their conclusions were that the magnitude of the magnetic field strength did not affect the amount of undercooling at which nucleation was initiated. While conclusive experimental results have been lacking to explain this effect of mechanical vibration on inducing nucleation, a number of theories have been proposed. Two of these theories are discussed below. 1) The Change in Melting:- Point Locally Due to the Change in Pressure (Clapeyron Equation). According to Vonnegut10 the most plausible explanation for the nucleation of a supercooled melt by cavitation is the effect of changing the melting point by a change in pressure. For materials where the volume decreases on solidification, an increase in pressure raises the melting point; for materials which expand on solidification, the melting point is raised for a decrease in pressure, i.e., rarefaction. Using the Clapeyron equation, the melting point of a metal can be calculated as a function of pressure. If it is assumed that the equation can also be used to calculate the temperature of nucleation of a supercooled melt as a function of pressure (i.e., the temperature of heterogeneous nucleation will increase with pressure at the same rate as the melting point), the amount of supercooling required for nucleation will be constant at all pressures as shown in Fig. 1. It is obvious that an isothermal change which results in an increase in melting point results in an equal increase in supercooling. This increase in supercooling may now be sufficient for nucleation. A pressure of 80,000 atm was calculated, using the Clapeyron equation, as the pressure required to increase the temperature of nucleation of nickel by 200°C. According to Lord Rayleigh,11 this very large pressure could be generated for a very brief period of time by the collapse of a cavity. This pressure wave is radiated in all directions from the collapsed cavity. If the temperature of the melt is slightly below its equilibrium melting temperature at atmospheric pressure, stable growth can follow; that is, once nucleation occurs, growth becomes the main driving force of the solidification process. This proposal has been extended to water which expands on freezing by assuming that nucleation occurs during rarefaction following the pressure pulse. This negative pressure pulse should follow immediately after the positive pressure pulse with its magnitude approaching the critical tensile strength of the liquid. The negative pressure developed during this period would raise the melting point of water and thus promote nucleation. Hunt and jackson12 have suggested this for water. Similarly, it could be postulated that bismuth which also expands on freezing could be nucleated during the negative pressure pulse. 2) Nucleation by a High-pressure Phase. An extension of the Clapeyron equation to systems where density decreased on freezing at atmosphere pressure has been proposed by Hickling.13 The phase diagram for water initially shows the well-known decrease in

Jan 1, 1969

By John Griffen

This paper records over twenty years&apos; experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years&apos; experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By John Griffen

This paper records over twenty years&apos; experience with the use of the Rheolaveur system in the United States, showing its ability to meet changing conditions caused by the dirtier mine output of present-day mechanical mining methods. Data are given on size ranges handled, number and capacity of units installed, maintenance experience, and operating results on two-product and three-product separations. IT is not the purpose of this paper to discuss the principles employed in the Rheolaveur system of fine-coal cleaning as these have been fully covered in the technical literature of the Institute."2 Rather, our purpose will be to record the highlights of over 20 years&apos; experience in the United States with Rheolaveur fine-coal launders, which will indicate their capabilities and costs of cleaning and their adaptability to meet the changing conditions caused by present-day mechanical mining methods. Rheolaveur fine-coal launders are used in the United States to treat a wide variety of size ranges of fine coal, the coarsest being 1/2 in. to 0 and the finest about 1/8 in. to 0. In Europe feeds as coarse as 5/8 in. and as small as 1/2 mm (28-mesh) to 0 are being cleaned. The size ranges usually handled in this country are 3/8 to 1/4-in. round to 0. Thirty five units with a combined hourly feed capacity of 3200 tph have been installed in the United States. Individual units are cleaning as little as 25 tph, while others are cleaning as much as 200 tph. Rheolaveur fine-coal launders offer cleaning units of high capacity and are outstanding in requiring a minimum of building space per ton of input. Experience has shown that maintenance costs are low. Several installations were made in the Pittsburgh district from 1928 to 1930, and since that time operation has been largely two shifts per day. Minor repairs to liner plates and Rheo boxes have been required during the intervening years, but no major replacement of launders was required until 1947 and 1948. At one of these plants, launders operated for 17 years before they were replaced and during that period over 26,000,000 tons of —5/16-in. coal were cleaned. The effect of the much dirtier raw coal produced by mechanical loading of the Pittsburgh seam is shown by the following tests. The data in table I summarize the performance of a Rheo fine-coal unit when cleaning hand-loaded raw coal. The —4-in. raw coal fed to the cleaning plant analyzed only 8.16 pct ash and 1.31 pct sulphur and contained 4.6 pct sink, 1.55 sp gr. Two years later, in 1945, mechanically-loaded coal from the same mine was cleaned in the same plant. The feed, —4 in., then analyzed 20.10 pct ash and 1.53 pct sulphur and contained 20.2 pct sink, 1.55 sp gr. The performance of the Rheo fine-coal unit when cleaning this coal is given in table 11. It will be noted that the +48-mesh coal is cleaned almost as thoroughly as in table I. The large amount of high-ash slimes produced from the dirtier feed are responsible for the higher ash of —--48-mesh washed coal in table 11. In this plant the Rheo fine-coal unit did not get the full load of refuse resulting from mechanical

Jan 1, 1951

By H. OSBORG

THE patent literature of alloys for the last two decades or so indicates that the number of liatents referring to smaller and smaller percentages of essential alloying constituents is on the increase, whereas the number of patents relating to considerable percentages of alloying additions becomes smaller. The claims regarding physical properties indicate consistently higher and more rigid values and show the possibility of modifying the physical properties by only slight changes in the percentage of minor constituents or by applying different methods - of treatment. Alloys of the same base have been made suitable for a number of different applications by intelligently and rigid- ly controlling their composition, melting, casting, working, and heat-treatments. Other trends indicated in the development of modern alloys are higher corrosion resistance, higher fatigue limits, and "stability" in volume.

Jan 1, 1937

By Roy A. Lindgren

Since the year 1643, when the first blast furnace in America for treating iron ore was built at Saugus, Mass., out of mica schist quarried in the neighboring district, the procurement of a suitable refractory for furnace lining has been a problem of concern to the operators of furnaces. The stacks built of mica schist continued to smelt iron ore until about 1836, when, according to F. H. Norton, the first firebrick were produced1, at Queens Run, Pa. Other writers speak of brick having been molded and burned in Massachusetts about the year 1834. In 1841, Andrew Russell began to produce medium refractory plastic clay brick near East Liverpool, Ohio, that were used for lining blast furnaces1. The well-known Kentucky clay-producing district was not opened up until the year 1871, but since then it has produced a large percentage of the linings for iron blast furnaces. While some strides have been made by the refractories industry during the 100 years that have passed since the first firebrick were produced, it has been only during the last two decades that any real progress has been made towards bettering the product, even though the method of production had improved. Perhaps the fault lies with the user of the brick rather than with the producer, for not sooner demanding a superior product. During the past 15 years the tonnage produced per lining has increased from 500,000 gross tons to 1,000,000 gross tons, and now some furnaces are producing 1,600,000 gross tons and better on a single lining. It is true that enlarged capacity of furnaces and improved practice have accounted for some of this increase in tonnage, nevertheless better quality in firebrick must be given credit for its share. However, we are not yet ready to say that we have reached a maximum life of furnace lining. We believe that a better product can be produced and that the refractories industries of America will, through their extensive research depart-

Jan 1, 1937