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For the year ending February, 1920 PRESIDENT HORACE V. WINCHELL MINNEAPOLIS, MINN. PAST PRESIDENTS PHILIP N. MOORE ST. LOUIS, Mo. SIDNEY J. JENNINGS NEW YORK, N. Y. FIRST VICE-PRESIDENT A. R. LEDOUX NEW YORK, N. Y. TREASURER GEORGE D. BARRON. . . . ' NEW YORK, N. Y SECRETARY BRADLEY STOUGHTON NEW YORK, N. Y

Jan 6, 1919

By Paul L. Hunter

Nearly 2000 people are employed at Kennecott's Chino Mines Division, which recently completed its 50th year as a leading copper producer. Chino operates an open pit mine at Santa Rita, N.M., and a reduction plant, including mill, power plant, and smelter, at Hurley, N.M. The property mines 66,000 tons of copper ore and materials per day. There have been safety and industrial health programs at Chino for a number of years, but in 1955 these procedures were reorganized on a more active basis. The following discussion evaluates the progress made during the 1955-1958 period.

Jan 6, 1960

By K. Huessener

This paper attempts a survey of the principles involved in the combustion of blast-furnace gas in boilers and stoves. I do not expect to be able to give much information which is actually new, since the laws and methods of calculation pertaining to this problem are contained in a number of handbooks; but, so far as I am aware, they have nowhere been treated exclusively with reference to this one problem of gas economy. The road which I follow is less that of strictly scientific research than that of the practical engineer, faced by operation problems which require a quick solution by easily accessible means without a prolonged search in scientific handbooks. For this purpose I give a number of tables and diagrams, as well as the results of many years1 practical working experience as combustion engineer, which I trust will prove useful to my readers. HlSTORICAL Considering the tremendous efforts made during the last 15 years to increase the efficiency of all kinds of power plants, it is highly surprising that until a comparatively very recent period, progress in the economical combustion of blast-furnace gas has been so slow, The explanation generally put forth, that fuel has been so cheap that it was hardly worth while to bother about this economy, will not hold good, for the greatest trouble was taken to improve the combustion of coal and to reduce the consumption of steam, The most plausible explanation appears to be the general impression, held as an accepted fact, that gas, especially hot, uncleaned, blast-furnace gas, was not a good boiler fuel. Many managers of high repute had tried unsuccessfully to improve the conditions of its use, and the dictum went forth that efficiencies of over 60 per cent., particularly with high loads, were impossible. This was the situation not only in the United States, with very cheap fuel, but also in Europe, where fuel prices are twice and three times as high.

Jan 1, 1916

PRESIDENT SIDNEY J. JENNINGS NEW YORK, N. Y. PAST PRESIDENTS L. D. RICKETTS NEW YORK, N. Y. PHILIP N. MOORE FIRST VICE-PRESIDENT C. W. GOODALE BUTTE, MONT. TREASURER GEORGE C. STONE, NEW YORK, N. Y. SECRETARY. EMERITUS ROSSITER W. RAYMOND, NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON, NEW YORK, N. Y.

Jan 5, 1918

By Arthur Gibson

Synopsis.-A new feature in coal mining, where the coal is to be conveyed from a high to a lower elevation and the topography of the country is such as to preclude surface haulage. The distance from the mine to the tipple inn this instance is about 3,000 ft. and the difference in elevation between the terminals 321 ft., of which 225 ft. occurs in the first 1,300 ft. from the loading terminal. A mine car was designed especially for this plant consisting of two tramway buckets on one pair of trucks, the buckets being lifted from the truck, attached to the train line automatically, and conveyed to the tipple. THROUGHOUT eastern Utah, in the Book Cliff range and the Wasatch plateau, the coal-bearing measures are admirably exposed, in the bold and for the most part bare escarpments. The geology is simple, the rocks being almost flat, or at the most dipping not more than 10 per cent. The coal beds occur in a formation, or group of strata, consisting of sandstone and shale, 600 to 800 ft. thick, which has been correlated with the Mesaverde formation of the Cretaceous. Above this occurs about 2,000 ft. of alternate layers of buff-colored sandstone and shale. Below the coal lie the Colorado shales, which are estimated to be about 1,500 ft. thick, below which are the red rocks of the Jura-Trias. As a rule the coal beds all clip in a northerly direction and under the high mountain. The canyons opening out into what is known as Castle valley all cut the coal beds, and the beds rising in the direction of the valley are found frequently high up on the escarpment, often as much as 1,000 ft. above the valley. The surface features of the region have an economic bearing on the availability of the coal. The valleys and gulches in the Colorado shale extend out from the coal areas across the edge of the plain, so that it is impracticable to build transportation lines parallel with the boundaries of the field within several miles of the coal land. It will be necessary, therefore, to build branch lines up to the coal lands from points on the Denver & Rio Grande Railroad traversing Castle valley.

Jan 10, 1914

By Scott Turner

AS IN OTHER lines of engineering, progress in mining was influenced during 1931 by the world-wide economic depression. Low-metal prices ? resulted in active efforts to reduce production costs of base-metal mines. Probably the chief factors in reducing costs were the greater efficiency of labor, due to better relations between management and workmen, to the scarcity of jobs, reduced wages in most base-metal mines, and to lower overhead expenses secured by drastic economies. The use of mechanical equipment continued to increase, due to more efficient designs and a better practical understanding of the applicability of such equipment to requirements. With one exception, there were no startling innovations in stoping methods, but there was notable progress in the better adaptation of mining methods. In the United States, no large new base-metal or silver mines came into production during the year. At a number of mines, output was curtailed, and operations were discontinued at several important producers. On the other hand world-wide interest in gold was stimulated, as evidenced by increased output from producing mines, the reopening of some old properties, and intensive prospecting for new deposits. Mining methods at several gold mines in the United States and Canada were modified or changed to obtain a higher extraction of ore or to meet the need for additional support of the mine workings at depth. Permanent record of improvements in mining practices has been made during the year by the Federal Government; the industry thus has available for its use more detailed information than ever before. This work was done through cooperation of the mining companies and their representatives with the United States Bureau of Mines, whereby many papers descriptive of practices and costs at individual mines have been published. The Bureau's program of describing methods and costs at the more important mines (those producing in excess of $100,000 annually) is nearly completed, but papers will be issued concerning such additional mines as may possess particular interest because of novel methods, exceptional conditions or operating results, and others to supplement reports already issued. In this way, an interesting record of the results of changing conditions, or innovations in methods, will be made available.

Jan 1, 1932

Jan 1, 1947

Jan 1, 1947

By G. K. Williams

The continuous process of lead refining as at present operated at the Port Piric plant of The Broken Hill Associated Smelters Proprietary Ltd. was a development from investigations conducted in connection with the ordinary Parkes desilverizing process. It is not proposed here to discuss the theoretical considerations involved in continuous desilverizing because this aspect has already been dealt with in previous publications.1 Continuous Softening Process For economical desilverization of lead bullion, the bullion must first be subjected to refining processes for the removal of such impurities as copper, sulphur, arsenic, antimony and tin. This refining is effected by first drossing to remove copper and sulphur and then preferentially oxidizing the arsenic, antimony and tin from the bullion in the softening process. The theoretical basis of the continuous softening process is illustrated by the portion of the Pb-Sb-O ternary equilibrium diagram shown in Fig. 1. Under conditions of complete liquefaction, metal of the composition of softened bullion is in equilibrium with an oxide slag containing about 9 per cent antimony. In practice, of course, arsenic also is always present, and while the 0Pb-Sb-As-O quaternary system has not been worked out in detail, it has been found that in softening drossed smelter bullion of the composition produced at Port Pirie, arsenic is always reduced to 0.001 per cent or less when the antimony in the bullion is reduced to a satisfactory figure. In the operation of the continuous softening process, a two-phase system of a bullion low in arsenic and antimony and an oxide slag relatively high in those components exists in equilibrium. To this system is added continuously a bullion containing 0.2 per cent As and 0.8 per

Jan 1, 1937

By R. A. Wattenbarger, H. J. Ramey

A systematic study has been made of the application of the real gas pseudo-pressure m(p) to short-time gas well testing. The m(p) function can be used in real gas flow problems to account for the variation of viscosity and z-factor with pressure. A mathematical model was solved numerically to generate solutions of various real gas flow problems. The effects of turbulence, formation damage and wellbore storage were included in the model. The analysis of simulated well tests showed that the interpretation methods normally used for liquid flow are generally accurate when the m(p) is used. For practical rates without turbulence, the solutions are not rate-sensifive, and the flow capacity.. kh and skin effect can be determined accurately from either a buildup or a drawdown test. However, the kh calculated from a drawdown test can be significantly low when turbulence is present. A case was simulated in which this error wa.s 36 percent. Turbulence does not affect the determination of kh from buildup tests. The proper determination of kh from p and pa buildup and drawdown plots is developed by analyzing their relationship to the m(p) method. A simple equation is given that can be used for long-range gas well performance forecasting. This expression is compared with a method pre-sended by Russell et al." Introduction The formation flow capacity and wellbore damage condition can be determined for liquid-producing tests by means of buildup and drawdown tests. These tests make use of short-time pressure transient data rather than stabilized flow tests that are often used for gas well testing. Tracy' presented a method of gas well testing that was based on ideal gas equations and utilized p2 buildup plots. This method was shown to be good for low pressure wells. MatthewsZ suggested plotting p and using an average slope and average gas properties. This method was more successful on high pressure wells. Al-Hussainy, Ramey and Crawfordh howed that variation in gas properties could be simplified by using the real gas pseu do-pressure m(p). Some cases showed that the use of the m(p) function Original manuscript received in Society of Petroleum Engineers office Aug. 8. 1967. Revised manuscript received June 11, 1968. Paper (SPE 1835) was presented at SPE 42nd Annual Fall Meeting held in Houston Tex., Oct. 1-4. 1967, and at SPE 4th Annual Eastern Regional ~eetin$ held in Pittsburgh. Pa., Nov. 2-3. 1967. 0 Copyright 1968 American Institute of Mining, Metallurgical, and Petroleum Engineers, Inc. 'References given at end of paper. provided an accurate method of interpreting buildup and drawdown tests. Because the work of Al-Hussainy et al. included only a few actual cases, there was a need to explore the use of the m(p) method for a greater variety of flow conditions. The occurrence of turbulent flow around the wellbore often is an important factor in gas well testing. Swift and Kiel,' and Carter et a1.' treated turbulence in gas well testing, but did not consider the variation of viscosity and z-factor with pressure. This paper presents the results of an investigation of the application of the m(p) method to buildup and drawdown testing. Vhe investigation included the effects of turbulence, formation damage and wellbore storage. Gas Flow Equations To formulate the mathematical model, many of the assumptions usually used in well testing theory are applied. The system has radial geometry with a closed outer boundary and is composed of a horizontal porous formation that has uniform and isotropic rock properties and uniform thickness. Allowance is made, however, for a radial region of reduced permeability near the wellbore. This region represents formation damage. The geometry of the system is shown in Fig. 1. Darcy's law does not always apply to gas flow. A more general expression is needed for non-Darcy, or "turbulent" REGION OF DAMAGED PERMEABILITY SEALED BOUNDARIES ClRCULAR WELL BORE Fig. 1—Radial flow model. AUGUST, BT7

Jan 1, 1969

By C. A. Meissner

The; following is a further discussion of the paper of James Qayley, " The Application of Dry-Air Blast to the Manufacture of Iron " (Trans., xxxv., 746), with special reference to his supplementary paper (Bi-Monthly Bulletin, No. 4, July, 1905, pp. 797 to 819), and to the paper of J. E. Johnson, Jr., " Notes on the Physical Action of the Blast Furnace " (Bi-Monthly Bulletin, No. 5, September, 1905, pp. 1111 to 1146). Mr. Gayley's paper has been discussed also by E. Windsor Richards, R. w. Raymond, E. H. Saniter, W. J. Poster, H. M. Howe, V, Pendred, B. H. Thwaite, James Gayley and Alexander Pourcel (Trans., xxxv., pp. 1022 to 1042), Joseph W. Richards (Bi-Monthly Bulletin, No. 4, July, 1905, pp. '703 to 718); T. W. Robinson (Bi-Monthly Bulletin, No. 4, July, 1905, pp. 788 to 796); C. B. Dudley, R. W. Raymond, J. E. Johnson, Jr., W. F. Mattes, James Gayley, David Baker, John Birkin-bine and F. E. Bachman (Bi-Monthly Bulletin, No. 5, Septem-

Jan 1, 1907

By AIME AIME

WIDESPREAD interest in the better utilization of coal is indicated by the attendance of over seventeen hundred men interested in the pro- cessing and utilization of coal and its by-products, at Pittsburgh, Nov. 15 to 18. The conference was truly international in scope and of a scientific character, and must be considered a real success. There were dele- gates--official and unofficial-from America, Belgium, Canada, Czechoslovakia, France, Germany, Great Britain, Japan, Mexico, and Sweden. Several Federal and 28 State departments had representatives, also many colleges and universities, and many leading industrial and mining corporations throughout the country. A feature of the sessions was that papers were mostly read in full and attentively listened to, although in some cases they were of a technical nature. The proceedings were opened by the president of Carnegie Institute of Technology, Thomas C. Baker, who spoke of the importance of the conference and felt that without foreign cooperation it could not have been held. There was no need to apologize for the movement of Carnegie Institute of Technology in the matter. Coal is commanding attention in every part of the world; so is power; and coal will remain the chief source of power for generations to come. The coal chemist is working with the engineer, and nobody can forecast what developments there will be in fuel technology. This conference may be the turning point of fuel technology in America. The development and application of these new processes may be of great benefit to Pittsburgh.

Jan 1, 1926

By L. D. Jaffe, R. K. Pitler

IT would clearly be advantageous to produce molten titanium, suitable for alloying and casting, directly from the relatively inexpensive tetra-chloride, without using a metallic reducing agent. Accordingly, a preliminary investigation has been made of the production of titanium by hydrogen reduction of the tetrachloride in an electric arc furnace. For the reaction: TiCl4(g) + 2H2(g) ? Ti(s) + 4HCl(g), Lockhart and his colleagues' have calculated equilibrium constants at 800" to 1500°K, based on tables of Brewer.' Extrapolating these data on the assumption that the heat of reaction varies linearly with temperature, and using Brewer's value for the heat of fusion, there is obtained for the reaction: TiCl4(g) + 2H2(g) ? Ti(l) + 4HCl(g) an equilibrium constant at 2100°K (100" above the melting point) Kp = (PHCl)4/pTlcl4 (Ph2)2 = 0.167 atm. If this is correct, starting with hydrogen saturated with TiC14 at 25°C, where the vapor pressure of the latter3 is 12.6 mm, and maintaining the overall pressure at 1 atm, 99 1/4 pct of the TiC14 would be reduced to Ti at equilibrium. An arc furnace designed for the melting of titanium and lined with water-cooled copper was used for the preliminary experimental check. Its inside diameter was 2 7/16 in. and height 6 in. The water-cooled electrode was tipped with 3/4 in. diam tung-sten. In the one run carried out, the furnace was charged with a starting batch of 293.35 g of scrap titanium,' analyzing 99.24 pct Ti. After the system was twice pumped out and flushed with argon to remove air, the arc was struck in argon and kept at 400 amp dc, electrode negative, for 1 min to melt the starting batch. Commercial tank hydrogen, dried by passing over silica gel and bubbled through technical grade TiC14 at room temperature, was then admitted to the furnace. The inlet tube directed this gas downward toward the arc and melt; the gas outlet was at the top of the furnace. The quantity of gas passed through was not measured accurately but is estimated to be about 50 liters per min at a few mm of mercury above atmospheric pressure. The arc was held in the H2-TiC14 mixture at 400 amp with little difficulty. After 10 min the arc burned through the copper lining above the melt, admitting water and ending the run. (This had happened frequently in straight melting runs under argon, and is not thought due to the atmosphere used.) In the furnace were found an ingot, several small pieces, and a little metallic powder, evidently formed when water entered the hot furnace. Some titanium also clung to the electrode. The portions were weighed and separately analyzed for titanium. Their total weight was 321.76 g of 94.82 pct overall purity. Much of the impurity consisted of tungsten broken off when the titanium was removed from the electrode. No detailed impurity analysis was made because the starting materials were impure and flooding the furnace introduced further contamination. Thus, the metal recovered contained 305.08 g Ti, as compared to 291.12 g in the starting batch. Titanium could not have been carried over from previous melting runs in the furnace, as the interior was sandblasted and the electrode ground clean before the run. Evidently at least 13.96 g of Ti were produced from TiCl4. Since this would require complete reduction of 430 liters of TiCl, gas at 12.6 mm of Hg, the efficiency of the process must have been rather high. Experimental work is continuing. No reason is seen why the process could not be scaled up to the size of a 30-ton steel-melting arc furnace. In a large unit there would be a large molten pool from which liquid metal could be withdrawn for alloying and casting. The problems of electrodes and power supply are similar to those arising in arc furnaces for melting titanium. Unreacted TiC14 and the HC1 in the exhaust gases could be recovered; unreacted H2 could be dried and re-used. The melt would be saturated with hydrogen, whose solubility in liquid titanium is unknown. Small amounts of hydrogen seem to have little effect on the properties of titanium.5 If the amount introduced was so large as to impair properties or cause casting difficulties, it could be reduced by holding the melt under an inert gas or by pumping on it, prior to casting. Acknowledgment The authors would like to thank I. Preble and M. Goldman for their valuable aid in the experimental work and the thermodynamic calculations, respectively. References 1R. J. Lockhart, J. J. Ward, M. J. Hussey, and J. W. Clegg: In unpublished report, "U. S. Air Force Project Rand. Titanium and Titanium Base Alloys." Battelle Memorial Institute (1949) 37-44. 2 L. Brewer et al: Chemistry and Metallurgy of Miscellaneous Materials—Thermodynamics. Ed. by L. L. Quill. pp 13-39, 60-192. (1950). McGraw-Hill Book Co. 3 K. Arii: Scientific Reports. Tohoku Imp. Univ. (1933) 22, 182. 4 R. S. Dean, J. R. Long, F. S. Wartman, and E. L. Anderson: Transactions AIME (1946) 166, 369-381; Metals Technology (Feb. 1946) TP 1961E. * R. I. Jaffee and I. E. Campbell: Transactions AIME (1949) 185, 646-654; Journal of Metals (Sept. 1949) TP 26813.

Jan 1, 1951

By M. E. V. Plummer, F. E. Beamish, J. C. Hole, J. M. Kavanagh

Prior researches have shown that the iron-copper-nickel content of platinum concentrates may be reduced by carbon to form a collecting alloy for the platinum metals in a manner exactly analogous to the formation of the lead fire assay button. The nature of the iron-copper-nickel alloys with the platinum metals has been investigated. Palladium and probably some of the other platinum metals form solid solutions with this alloy. ONE of the greatest known reserves of the platinum metals is to be found in the nickel, copper, iron sulphide deposits of Northern Ontario. The very large tonnage of copper-nickel ore processed makes the Sudbury district a significant world source of the platinum metals, the latter being recovered as byproducts in the electrolytic refining process for copper and nickel. Chalcopyrite, pentlandite, and pyrrhotite are carriers of the platinum metals;' these minerals contain from 0.03 to 2.0 ppm of the noble metals. The analysis of the above ores for the platinum metals necessitates an initial concentration of the minute quantities present in the ore. The method usually employed involves the extraction of the noble metals by molten lead with or without small proportions of silver. The lead is then separated from the precious metals by cupellation, a process which involves absorption by the cupel and volatilization of lead oxide, leaving the precious metals collected in a silver bead on the surface of the cupel. The collection by lead is probably achieved by a process of dissolution followed by some degree of segregation in the lead alloy since the cold button is seldom uniform in composition. With silver-lead alloys this process of liquation has been used to re- cover pure lead since, during the cooling period, crystals of pure lead separate and silver becomes concentrated toward the center of the solid.' Platinum and gold liquate in a manner similar to silver which may account for the successful recovery of platinum by fire assay even though it is considered insoluble in a solid lead medium. 3 In the case of iridium the liquation process may be a disadvantage in that the iridium may appear in the outer areas of the button and thus be lost mechanically during the handling processes. In any case the deficient alloying characteristics of lead for the platinum metals encouraged attempts to find an improved collecting alloy, and by analogy with aqueous extractions, preferably one in which solid solutions would be readily formed. Since, beyond a certain minimum solute to solvent ratio, the formation of a solid solution is accompanied by a lattice distortion, one would expect some relationship between the atomic diameters of any two metals and their tendency to form a solid solution. In one survey5 the hypothesis was put forward that when the atomic diameter of the solute differs by more than about 15 pct from the solvent, the "size factor" is unfavorable and the formation of a solid solution is restricted. On the other hand, when the atomic diameters are within this limit the size factor is favorable and solid solutions are readily formed, e.g. the transition metals of Group VIII having similar radii form solid solutions over wide ranges of composition.6 Even with a favorable size factor, elements having different crystal structures cannot form a continuous series of solid solutions. Cell dimensions alter as the concentration of the solute increases, causing a

Jan 1, 1962

By C Macquarie, Maurice Cook

Normally the grain size of cold-rolled and annealed copper sheet is of the order of 0.02 to 0.06 mm., and 0.1 mm., for example, would, for many purposes, be regarded as undesirably large. The occurrence of very large grain sizes in a few isolated cases of copper sheet led the authors to inquire into the causes giving rise to such unusual results. During the course of this work a very large number of experiments have been carried out to determine the effect of varying the conditions of the final and penultimate rolling and annealing operations, as well as of earlier rolling and annealing treatments. A correspondingly large number of observations have been made, but within the space of one paper it is possible only to outline relatively briefly what appear to be the more important of these experiments and observations. The presentation of the results does not follow chronologically the order in which the experiments were carried out and they have been rearranged for the sake of simplicity. In recent years a considerable amount of work has been done in connection with the recrystallization and grain growth in metals arid alloy. that take place on annealing, and a number of generalizations have beell reached. It is known that the crystal size in wrought material in the annealed condition depends on a number of factors, including the time and temperature of the annealing operations and the nature and amount of cold-work to which the material has been subjected. Results. however that are not covered by these generalizations can be obtained, which, in some instances, may even appear to be contradictory to them. Xccortling to these generalizations gradual changes in grain size are effected by progressive variation of the factors involved in annealing or cold-working. The work recorded in the present paper, however, relates to the development of grain sizes which are far beyond those normally encountered in ordinary annealing operations and far in excess of grain sizes which,

Jan 1, 1939

By C Macquarie, Maurice Cook

Normally the grain size of cold-rolled and annealed copper sheet is of the order of 0.02 to 0.06 mm., and 0.1 mm., for example, would, for many purposes, be regarded as undesirably large. The occurrence of very large grain sizes in a few isolated cases of copper sheet led the authors to inquire into the causes giving rise to such unusual results. During the course of this work a very large number of experiments have been carried out to determine the effect of varying the conditions of the final and penultimate rolling and annealing operations, as well as of earlier rolling and annealing treatments. A correspondingly large number of observations have been made, but within the space of one paper it is possible only to outline relatively briefly what appear to be the more important of these experiments and observations. The presentation of the results does not follow chronologically the order in which the experiments were carried out and they have been rearranged for the sake of simplicity. In recent years a considerable amount of work has been done in connection with the recrystallization and grain growth in metals arid alloy. that take place on annealing, and a number of generalizations have beell reached. It is known that the crystal size in wrought material in the annealed condition depends on a number of factors, including the time and temperature of the annealing operations and the nature and amount of cold-work to which the material has been subjected. Results. however that are not covered by these generalizations can be obtained, which, in some instances, may even appear to be contradictory to them. Xccortling to these generalizations gradual changes in grain size are effected by progressive variation of the factors involved in annealing or cold-working. The work recorded in the present paper, however, relates to the development of grain sizes which are far beyond those normally encountered in ordinary annealing operations and far in excess of grain sizes which,

Jan 1, 1939

By Frank C. Roberts

Progress in the facilities for handling mining products has been largely superinduced by the necessities of commercial economy ren dered requisite in order to meet the demand of competition. So rapid has been the depreciation in the value of mineral products, primarily due to the disproportionate increase of output over con sumption, that it may be justly claimed that mining, when considered as a commercial success, depends largely upon the ease and cheap ness by which the products are brought to the surface of the ground. In all methods of performing this operation, wire rope enters as an important factor; and the object of the present sketch will be to explain, as fully as is consistent with the space allotted, the various • adaptations of wire rope employed in placing coal and ores within the reach of our overground systems of transportation. These adaptations will be considered in the following order: I. Hoists. II. Inclined planes. a. Engine planes. b. Gravity planes. c. Aerial planes. III. Haulage. a. Tail-rope system. b. Counter-rope system. c. Endless-rope system. IV. Hoists. The term " hoist" is applied to the system whereby the mine product is elevated from the mine-level to the surface through ver-

Jan 1, 1888

By T. F. McCormick

THIS year, 1950, is the golden anniversary of the construction and operation in this country of a tube mill for the sole purpose of fabricating aluminum alloy tubing. For a short period prior to the beginning of this mill, aluminum alloy tubing was obtained from foreign sources or produced in mills fabricating other metals. Indeed, it was a very short period because aluminum itself was a newcomer in the field of metals. The first sheet rolling mill had been in operation just eight years and a rod rolling mill for the production of wire and cable had been functioning one year. The first tube mill for aluminum was built by the Aluminum Co. of America in May 1900 and started production with four draw benches and five men, including the mill superintendent. The building itself was a lean-to attached to a sheet-rolling mill at New Kensington, Pa. The cost of equipment, including the draw benches, dies, mandrels, and other tools, was about $5000. This figure seems absurd in the present day but nevertheless it was questioned at that time whether or not such a large sum should be spent to embark on a new and uncertain venture. The initial production system for the new mill consisted of casting round hollow ingots in a tilting-type iron mold, reducing the ingots to bloom size outside the plant, and finally drawing to size on the draw benches. This method of producing aluminum alloy tubing survived just two years, as blooms of satisfactory quality were not obtained consistently. It was superseded by the cupping method in which a 24-in. diam circle cut from a rolled plate was formed into a tube bloom using first a small cupping press and then a push bench. It was stated that tubing so produced was limited to a maximum diameter of 2 3/4 in. and to wall thicknesses varying from 12 to 24 gage. However, for the first time all the operations were performed at one plant and the result was a quality product. Later larger cupping presses and push benches were obtained and the method continued to be the principal one for aluminum alloy tubing for almost a quarter of a century. During the interim the use of a Alan Mannesmann -type billet piercer was explored and it was used to some extent, but failed to produce consistently a smooth inside surface suitable for drawing with aluminum alloys. Experiments started during the latter stages of World War I in the production of tube blooms with a hydraulic extrusion press, led to the use of this type of equipment to make tube blooms for further drawing during a change-over period extending from 1925 until 1930. At this time the cupping method was discarded completely. From these humble beginnings, the aluminum alloy tubing business expanded until well over six million pounds were produced in a single month in the United States during World War II. The major portion of this production was in strong alloys which did not appear in the tubing picture until 1022, when 17S alloy tubing was introduced for aircraft construction.

Jan 1, 1951

By F. E. Smith, David R. Mitchell, B. M. Bird

DENSE-MEDIA processes utilize the familiar laboratory float-and- sink procedure on a commercial scale. Just as wood chips float on water and sand sinks, ,so coal floats and refuse sinks when placed in a medium intermediate in specific gravity between the coal and refuse. Liquids intermediate in specific gravity between coal and refuse used in commercial plants are of four types: (1) Organic liquids of high specific gravity-halogenated hydrocarbons; (2) dissolved, salts in water; (3) pseudo liquids obtained by suspending solids, such as sand, in water, and (4) fluidizing a mass of sand or other solids by air currents. Those most widely used for cleaning coal are solutions of calcium chloride in water and a suspended-sand pseudo liquid. Laboratory float-and-sink testing is an intermittent operation during which every opportunity is given for particles lighter than the medium to float and particles heavier to sink. Commercial processes are continuous, vertical and horizontal currents of the separating medium occur as either part of the process or unavoidably produced by the introduction of feed and withdrawal of products, so that the exact separation of the laboratory technique is not duplicated by commercial processes. However, separations close to the theoretical are obtained in commercial machines by careful technical control of the process. Most machines in commercial installations are crowded to get as much out of the equipment as possible, so that careful technical control is necessary to get the recoveries and grade of product desired. Sizes usually cleaned are plus 3/16-in. or plus 1/8-in. in the United States and plus 1/16-in. in Great Britain: Coal as small as 1 mm. has been cleaned in some experimental installations. The separation of coal from refuse in sizes finer than about in. is made difficult by viscosity' effects and hence at the present time cleaning is confined to the coarser sizes. Technical advances in the art of cleaning coal by dense-media

Jan 1, 1943

By J. L. Huitt, S. R. Darin

The use of a partial monolayer of propping agent to obtain a high flow capacity for a hydraulically induced fracture is discussed. From the results of laboratory work it was shown that a modified form of the Kozeny-Carman relation could be used to describe the flow in the partial monolayer propped fracture. With equations presented in the paper, the density pattern of the propping agent (number of particles per unit of fracture surface) that results in the maximum flow capacity for the fracture can be determined. The maxitnum flow capacity obtained with a partial monolayer is often an order of magnitude greater than the flow capacity obtained in greater width fractures containing multilayers of the propping agent. INTRODUCTION One of the predominant factors controlling the success of a hydraulic fracturing operation is the propping of the fracture. The trend in hydraulic fracturing recently has been an increase in the ratio of propping agent-to-fluid. The primary purpose of this increased ratio is to sustain a propped fracture of greater width. It was shown in a recent study' that for some formations low concentrations of propping agent would result in "closing" or "healing" of the fracture. In some cases, in an effort to insure sufficient propping agent concentration, the fracturing operation is designed to obtain a pack of the propping agent in the fracture. The placing of a pack of propping agent in the fracture provides a fracture of maximum width; however, the width alone does not control the flow capacity of the fracture. The flow capacity is dependent on the permeability of the fracture, as well as the fracture width. Thus, increasing the permeability to obtain larger flow capacities is as important as obtaining a greater width fracture. By increasing the concentration and particle diameter of the propping agent, fractures of sufficient flow capacities are obtainable for effective well stimulation in a number of formations. However, in some of the recovery techniques, the fracture capacity obtained by commonly used practices in fracturing is inadequate. One such recovery technique is gravity drainage via a horizontal fracture placed near the base of the productive zone. For this recovery technique, the fracture capacity desired is usually one to two orders of magnitude greater than the fracture capacity normally obtained in fractures propped with the size of sand commonly used (20-40 mesh). By the use of particles of larger diameter, the permeability of a pack2 is increased; however, even a multilayer pack results in a fracture flow capacity lower than that desired for some formations. For example, a fracture of 1/4-in. width, packed with 8-12 mesh sand of about 2,000 darcies permeability, would result in a fracture flow capacity (permeability X fracture width) of only 40,000 md-ft. If the fracture flow capacity needed is an order of magnitude greater than that cited in the example, an approach different from the formation of a pack in the fracture is needed. One approach is a partial monolayer of large-diameter propping agent. In working with this type of fracture system, it was found that the flow problems were unique to the system; therefore, the study discussed in this paper was made. THEORY AND RELATED STUDIES Previous studies, although related to that described in this paper, have not pertained directly to the flow in a fracture propped with a partial monolayer of propping agent. However, studies such as flow between parallel plates,' flow in packed columns,' fracture flow capacity' and flow in simulated fractures reduce the scope of this flow problem considerably. A fracture propped with a partial monolayer of dropping agent may be considered to be between two extremes — between an open fracture and a bed of particles packed In a conduit of fracture geometry. Thus, it seems appropriate to review the extreme cases.