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By T. A. Klinefelter

Lightweight concrete aggregates are materials weighing less than the usual aggregates of sand, gravel, and crushed rock. Concretes made with sand and gravel or crushed rock weigh 145 to 150 lb per cu ft, compared with 50 to 90 lb per cu ft for concretes made with lightweight aggregate. The ultra-lightweight concretes weigh much less than 50 lb per cu ft. Lightweight aggregates may be natural or manufactured materials, or byproducts from other commercial operations. Natural materials include pumice and pumicite, tuffs, breccia, scoria, diatomite, and vegetable products, such as peat, straw, and sawdust. Manufactured materials include expanded perlite, vermiculite, slag, clay, shale, and slate. Byproducts include cinders, aircooled slags, and coke breeze. Concretes made with lightweight aggregate were used first in construction of commercial buildings, later in homes. Present usage is general in multistory and other building construction, such as bridge decks, barges, jet-plane runways, piers and other structures. The industry has grown remarkably in the last decade. General Considerations The particular use to which the concrete may be put determines the importance of the various properties of the aggregates employed. Any given type of aggregate, however, should be reasonably uniform in composition and properties. It should be suitably graded and maintain proper weights, insulation values, and adequate strength to prevent breakdowns in mixing. It should bond properly with the cement, be inert to chemical reactions with cement and steel, and resist weathering, moisture, insects, and fungi. Some properties of any given type of lightweight aggregate from different suppliers may vary considerably because of differences in the source of raw material and the method of treatment. This statement applies particularly to expanded clays, shales and slates. All lightweight concretes have higher absorption of moisture and lower thermal conductivity, lower compressive strength, as well as lower unit weight as compared with standard concretes made with sand, gravel or crushed rock. Excepting cinders, the cost of lightweight aggregates is usually higher. Grading and unit weight requirements are given in ASTM standards: C 330-53T, Lightweight Aggregates for Structural Concrete; C 331-53T, Lightweight Aggregates for Concrete Masonry Units; C 332-54T, Lightweight Aggregates for Insulating Concrete. The production of masonry units shows the growth of the use and types of lightweight aggregates. In 1947 the total number of lightweight block was 460,000,000 units in 8 by 8 by 16-in. equivalents. By 1954 this had increased to 950,000,000 units. Of the 1954 production, 27 pct was of manufactured types, 43 pct of cinders, and 30 pct of remaining types. Raw Materials Used as Lightweight Aggregates It was stated above that a variety of products both natural and manufactured are used as lightweight aggregates. It is easy to understand that a bulky material of low unit value cannot be transported far and still remain a low cost product. Hence the product used in a particular market depends upon the availability of the possible raw materials. Thus it is

Jan 1, 1960

By A. J. Lyon, J. B. Johnson, Samuel Daniels

The wrought alloys of aluminum with small amounts of copper and of magnesium have, with the development of the automotive and aircraft industries, sprung into prominence through the medium of duralumin; but the cast alloys remained in the background because their possibilities in the heat-treated condition were either not recognized or not fully investigated. In 1919, the Bureau of Standards published a preliminary survey1 of this field, which pointed out that, in general, the ultimate strength and the percentage of elongation of the alloys of aluminum with and without magnesium could be improved by so simple a treatment as heating at 932' F. (500" C.) for 2 hr., cooling in air, and aging at room temperature for several days. British investigators, too, working with chill-cast material, found that" considerable improvement in the mechanical properties of copper-aluminum alloys may be obtained from suitable heat treatment applied to castings."' Jeffries and Gibson8 made the first contribution to the literature dealing exclusively with the heat treatment of sand-cast aluminum-base alloys, describing the effect of small additions or of combinations of copper, magnesium, and iron on the response to a short heat treatment comprising heating at 932" F. in a nitrate bath for 1 hr., quenching in fish oil, and reheating at 300' F. for for 1 hr. Their results disclosed the facts that the ultimate strength and the percentage of elongation of sound alloys were usually, and quite uniformly, improved by heat treatment; that with the iron content ranging between 0.36 and 0.88 per cent. and the silicon content between 0.24 and 0.60 per cent. the effect of increasing the copper content (Tables 10

Jan 1, 1925

By AIME AIME

THE outstanding event of the past month has been the conference in Washington of the representatives of about seventy-five of the leading national, regional and local engineering and allied technical organizations." Editorial-July Journal-The American Society of Mechanical Engineers. ?Prominent engineers from all parts of the country, representing more than 75 engineering societies, with an aggregate membership of more than 125,000, participated in the organization.? Editorial-July Journal-American Institute of Mining and Metallurgical Engineers. ?Civil, mechanical, electrical, mining, heating, sanitary, chemical, agricultural, marine, refrigerating, radio, fire protection, automotive, industrial and military engineers, as well as some architects and geologists assembled at Washington, D. C. to consider the advisability of organizing a national federation capable of bringing the united efforts of the entire membership to a focus on any problem of mutual interest." American Architect, June 23, 1920. ?"the object sought places it above any engineering gathering ever held." Editorial-American Machinist, June 17, 1920. "The youngest. profession has voluntarily accepted its responsibility in human affairs and is now reaching forward to a career of essential service." "This is .unmistakably the meaning of the Organizing Conference " Editorial-Industrial Management, July; 1920.

Jan 1, 1920

By Frederick H. McDowell

Notwithstanding the rapidly increasing use of copper, due to the extension of its applications within the last few years, the fact of its continued steady decline in price stands prominently forward. Nor have we far to look for the cause. The opening up of the vast copper-deposits of our Western States and territories is the metallurgical event of the last five years. We must not forget, in looking at the new field, the changed conditions under which the metallurgist must work, as compared with the previous history of copper-smelting. Until quite recently, the copper-plants of the country were few in number; and the concentration of industry implied in this circumstance, allowed many practices which must be considered as inapplicable to the conditions offered at the many mines which have of late become such important factors in our copper-production. Thus the old style of smelting the ore for matte in brick furnaces, involving great expense and time in their construction, while deservedly losing repute even in its old strongholds, in comparison with water-jacketed smelters, is still in some cases adhered to. But to put up such furnaces at Western mines would be out of the question. The large and continual need of brick for relining and repairs acts as an effectual bar, when such furnaces come into competition with the light, portable, and easily managed water-jacketed smelters of improved types. It is owing to the present unusual activity, manifested in the opening up of new copper-properties and the erection of plants for working them, that any improvement in furnace-construction is of especial interest and importance. The method in which processes are copied without variation in the rapid growth of metallurgical interests, has already been strikingly exemplified in the history of

Jan 1, 1885

By A. M. Gaudin, Olav Mellgren, P. L. De Bruyn

In commonly used to prepare the surface of the mineral to be floated so that attachment to air takes place. The quantity of agent required to make the mineral hydrophobic is usually very small, of the order of 0.1 to 0.25 lb per ton of mineral. Details of the mechanism of pyrite collection are for the most part unsettled. Adsorption of collector has long been believed to involve an ion exchange mechanism as demonstrated for galena' and for chalcocite. In the work on chalcocite it was also demonstrated that a film of xanthate radicals unleachable in solvents that dissolve alkali xanthates, copper xanthate, or dixanthogen was formed at the surface of the mineral. The unleachable product increased with increasing addition of xanthate up to a maximum corresponding to an oriented monolayer of xanthate radicals. Pyrite is extremely floatable with xanthate if its surface is fresh,,' but the floatability decreases rapidly as oxide coatings increase in abundance. Pyrite shows zero contact angle when in contact with ethyl xanthate solution at pH higher than about 10.5; at neutrality. a contact angle of 60° is obtained at a reagent concentration of 25 mg per liter. Alkali sulfides and cyanides are pyrite depressants.

Jan 1, 1956

By C. F. Seaman, R. G. Clay

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By R. G. Clay, C. F. Seaman

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By C. E. Gregor

Current modern trends in mining show that hoisting shafts are being expected to fulfill an important ventilation function. However, where rigid guides and supporting structures are mandatory, ventilating power costs will be high, unless the transverse structures (shaft sets) are streamlined. In the complex configurations of hoisting shaft furniture patterns, either for original design purposes, or for the adjustment of an existing shaft, a useful degree of aerodynamic optimization, by reduction of overall resistance, can only be effectively evaluated by model studies in a special purpose wind tunnel, as described herein. The author has shown that, in one particular case, an annual power cost saving of more than $50,000 was indicated, mainly by the selection of buntons and dividers of improved aerodynamic profile. Mining companies are now realizing the advantages that can accrue through proper attention to the aerodynamic aspects of mine shaft design. With progressive demands of civilization for an ever greater supply of coal and metals, and with greater depths of working, higher rates of extraction, advanced degrees of mechanization, and improved standards of environmental health and work comfort, shaft design is becoming an extremely important aspect of mining operations. In addition, because the cost of sinking, equipping, maintaining, and operating a mine shaft is continually increasing, there is a greater incentive to use a shaft for the maximum duty, consistent with economy. The current natural tendency is therefore to sink multi-purpose shafts, to install as much equipment as is safely and effectively possible to operate; and then, over and above all these functions, to use the shaft to its maximum possible ventilating capacity, as a major, or auxiliary, airway. Clearly, this latter additional function can only be employed at a significant cost, since the most economic operating airways are smooth-lined and devoid of internal rigid furniture. It is true that hoisting functions can be carried out in such shafts if rope guides are used, but only under restricted conditions. Where a manway must be provided, fixed guides and supporting structures are mandatory, and ventilation costs become accentuated. It is generally recognized that the prime element of cost in ventilating a mine can be represented by the power required to drive the fan unit. It is also known that the largest component of total mine resistance is that of the downcast shaft, where this shaft is equipped with rigid guides for hoisting purposes. The pressure loss varies directly as the resistance of the shaft, in accordance with a modification of the Atkinson equation p = RQ2/3600. Furthermore, the air horsepower drawn is given by the equation AHP = (p x 0/33000. Therefore, the power required to drive the fan unit varies as the resistance of the shaft and the cube of the air quantity. Clearly, if Q is a specified minimum, then the only way to minimize the pressure loss (and the power costs) is to reduce the shaft resistance R. (See Table I for a list of symbols.) The resistance to airflow of a shaft equipped with rigid guides and supporting structural sets consists of two components which, for the sake of analytical simplicity, can be regarded as additive. 1) Frictional resistance, represented by the viscous frictional forces within the boundary layer, as air flows along the surfaces of the walls and the periphery of the vertical elements (guides, pipes, cables, ropes, etc.) of the furniture. The friction loss can be readily calculated by the Atkinson equation, after selection of a suitable value for the friction factor K from the McElroy compilation.' KxLxO xQ2 5.2 xA3 Friction loss can be minimized by lining the shaft walls (particularly if of circular or elliptical cross-section) with a smooth continuous lining of monolithic concrete.

Jan 1, 1968

FOR this Famous Mining Enterprise issue MINING ENGINEERING selected the company that started the zinc in dustry in the United States. The New Jersey Zinc Co. has been a supplier of zinc products to the nation's industries for more than 104 years. Its Horse Head trade mark is famous throughout the world.

Jan 12, 1953

By Y. M. Chang, C. H. Pitt

A jet slurry apparatus has been constructed to test corrosive/erosive wear properties of grinding ball metal. Plain high carbon steel heat treated to a hardness of about 60 Rc was used. Slurries of quartz and chalcopyrite ore concentrate were used as the jetting media. Weight loss due to pure corrosion was determined by the linear polarization technique. Weight loss due to pure erosion was determined by jetting while cathodically protecting the sample. By comparing the total weight loss with the sum of the weight losses from pure corrosion and pure erosion, it was found that there was a synergistic effect between corrosion and erosive wear. Corrosion rate, erosion rate, and total erosion/corrosion rate are given as a function of pH and mineral type.

Jan 1, 1986

By A. A. Venghiattis

A rational approach to the selection of the appropriate perforator to use in each specific zone of an oil well is presented. The criteria presently in use for this choice bear little resemblance with actual down-hole condilions. These environmental conditions affect the elastic properties of rocks. One of these elastic properties, acoustic velocity, is suggested as the leading parameter to adopt for the choice of a perforator because, being currently measured in the natural location of the formation, it takes into account all of the effects of compaction, saturation, temperature, etc., which are overlooked in the laboratory. Equations and curves in relation with this suggestion are given to allow the prediction of the depth of perforation of bullets and shaped charges when an acoustic log has been run in the zone to be perforated. INTRODUCTION When an oil company has to decide on the perforator to choose for a completion job, I wonder if it is really understood that, to date, there is no rational way of selecting the right perforator on the basis of what it will do down-hole. This situation stems from the fact that the many varieties of existing perforators, bullets or shaped charges, are promoted on the basis of their performance in the laboratory, but very little is said on how this performance will be affected by subsurface conditions such as the combination of high overburden pressure and high temperature, for example. The purpose of this paper is to show the limitations of the existing ways of evaluating the performance of perforators, to show that performances obtained in laboratories cannot be extended to down-hole conditions because the elastic properties of rocks are affected by these conditions and, finally, to suggest and justify the use of the acoustic velocity of rocks, as the parameter to utilize for the anticipation of the performance of a perforator in true down-hole environment. EVALUATING THE PERFORMANCE OF A PERFORATOR It is natural, of course, to judge the performance of a perforator from the size of the hole it makes in a predetermined target. Considering that the ultimate target for an oilwell perforator is the oil-bearing formation preceded in most cases by a layer of cement and by the wall of a steel casing, the difficulties begin with the choice of an adequate experimental target material. For obvious reasons of convenience, the first choice that came to the mind of perforator designers was mild steel. This is a reasonable choice for the comparison of two perforators in first approximation. Mild steel is commercially available in a rather consistent state and quality, and is comparatively inexpensive. The trouble with mild steel is that it represents a yardstick very much contracted; minute variations in depth of penetration or hole diameter and shape may be significant though difficult to measure. The penetration of projectiles in steel being a function of the Brinell hardness of the steel (Gabeaud, O'Neill, Grun-wood, Poboril, et al), it is often difficult to decide whether to attribute a small difference in penetration to a variation on the target hardness or to an actual variation on the efficiency of the projectile. Another target material which has been widely used for testing the efficiency of bullets or shaped charges in an effort to represent a formation—a mineral target as opposed to an all-steel target—is cement cast in steel containers. This type of target, although offering a larger scale for measuring penetrations, proved so unreliable because of its poor repeatability that it had to be abandoned by most designers. The drawbacks of these target materials, and particularly their complete lack of similarity with an oil-bearing formation, became so evident that a more realistic target arrangement was sought until a tacit agreement was reached between customers and designers of oilwell perforators on a testing target of the type shown on Fig. 1. This became almost a necessity about seven years ago because of the introduction of a new parameter in the evaluation of the efficiency of a perforator, the well flow index (WFI). The WFI is the ratio (under predetermined and constant conditions of ambiance, pressure and temperature) of the permeability to a ceitain grade of kerosene of the target core (usually Berea sandstone) after verforation. to its vermeabilitv before perforation. The value of this index ;or the present state if the perforation technique varies from 0 to 2.5, the good perforators presently available rating somewhere around 2.0 and the poor ones around 0.8, There is no doubt that, to date, the WFI type of test is by far the most significant one for comparing perforators. It is obvious that a demonstration of a perforator

By H. S. McQueen

The insoluble-residue method for the examination and correlation of limestones and dolomites, or other sedimentary rocks containing calcium and magnesium carbonates, originated and was developed in the laboratory of the Missouri Geological Survey. The numerous and ever-increasing problems in the field of applied geology, and the need for a tool or aid in the solution of them, led to experimental work with samples collected from deep wells, in the early part of 1924. The success attained, almost at the outset, with this method was so great that the method has now been developed from the state of the experimental to the realm of large-scale, every-day routine. In a previous paper the writer1 described the procedure in detail and the characteristics of the residues obtained from certain Paleozoic formations in Missouri. The purpose of the present paper is to discuss the economic application of the insoluble-residue method. Developmental History Within the Ozark region of southern Missouri the geologic column consists of Lower Ordovician (Canadian of Ulrich) and Upper Cambrian formations, which are composed mainly of dolomite with subordinate amounts of sandstone, shale and limestone. Chert is found to some extent in all except one of the dolomite formations. The stratigraphy of the region is complex because the dolomites have lithologic similarities, fossils are few and often poorly preserved, and identifiable horizons within the comparatively thick formations are not common. Although considerable stratigraphic work of an areal nature had been done for a period of many years by the Missouri Geological Survey, the detailed geologic succession had not, until a few years ago, been systematically worked out. The problems of subsurface stratigraphy were even more complex. This region constitutes the main ground-water province of the State of Missouri and within it wells of varying depths are drilled in order to obtain supplies of ground water. In the work of the State Geological Survey in connection with the drilling of water wells for public supplies * Published with the permission of the Director, Missouri Geological Survey and Water Resources. Manuscript received at the office of the Institute Feb. 20, 1936. Assistant State Geologist of Missouri; Missouri Geological Survey, Rolla, Mo. References are at the end of the paper.

Jan 1, 1937

By Charles Jones

Discussion of the paper of CHARLES COLCOCK JONES, presented at the San Francisco meeting, September, 1915, and printed in Bulletin No. 105, September, 1915, pp. 1887 to 1898. D. A. LYON, Salt Lake City, Utah.-For some years I have been interested in the iron-ore deposits of the Pacific Coast, and I feel very much indebted to Mr. Jones for the information he has given us on this subject. I think that we are also very much indebted to him for having presented this paper, as he has given us reliable data in shape of facts and figures, which is just what we need in considering whether or not it would be commercially feasible to attempt to establish an iron and steel industry on the Pacific Coast. Personally, due to lack of market, I do not think it is a commercially feasible proposition at the present time. Mr. Jones states that pig iron can be produced on this coast as cheaply as it can be produced in Pittsburgh. Even. so, what can be done, with it after it is produced? To begin with, iron produced on the. Pacific Coast has to compete with iron produced in China. Last year I had occasion to investigate the subject of the market for iron and, steel on the Pacific slope, and found that it was possible for the producers of Chinese iron to sell it for much less than the price for which iron can be produced in the West, and also for less than eastern pig iron is now sold for, even when shipped through the Panama Canal. I was informed that the agents for Chinese iron had instructions to sell it for $1 or $2 less per ton than that asked for other irons on the market. Of course, I realize that the competition of Chinese iron may not be important, but even if it is not, and even though pig iron ,can be produced at the cost given by Mr. Jones in his paper, I do not see any immediate chance of its being so produced, due to lack of market. C. C. JONES, Los Angeles, Cal.-Mr. Lyon is right. It is a fact that Chinese pig is being sold in San Francisco on some large contracts at $18.50 a ton, while $19 is the usual price. We ought to produce iron. on this coast at $14. Personally I do not think Chinese iron is ever going to be an important factor. It is a bugbear out here now. You are not going to have much Chinese pig iron in .this country. Our little friend Japan is going to absorb everything it can. The price of Chinese pig iron at $18 is about as low as it will ever get. If we can show you that we can produce this iron here at the same price as at Pittsburgh, I. think you will have a good opportunity to compete with anything China or India can ever do. Costs there are very low now. They are going to come up by leaps and bounds in the next few years. The labor cost over there may be very low now, but I. have seen some men from there quite recently

Jan 12, 1915

By Carl Trischka

During fifty-three years of mining operations in the Warren mining district, the mineralization has been found to cover an area roughly 2 miles long by 155 miles wide. Ore extraction from the richer portions of this area has been closely followed by surface subsidence. At the western end of the district, where the ore outcrops at the surface and oxidation is more complete over a wide area, intense surface disturbance is visible in the form of closely spaced open cracks and escarpments. As work has progressed toward the east, more unoxidized ore-bodies have been found and mining has been at increasing depth, for which reasons surface subsidence is less intense and subsidence blocks become larger and in general are bounded by open cracks, which persist over long distances. There may also be traced on the surface what appears to be a boundary crack, running in a northwest-southeast direction, marking the present southerly limit of subsidence of the district as a whole. It is not completely developed, and has not as yet appeared over the very deep mining operations in the extreme easterly portion of the district. Check surveys show that vertical and horizontal dislocation is still taking place. All of the various mining operations, including open-pit work in porphyry ore and underground mining of limestone replacement and porphyry ores, have had their influence in disturbing the surface. As these influences frequently overlap, and because the surface disturbance is continuous over practically all ore occurrences, it is virtually impossible to choose an area sufficiently isolated to ascribe the amount of subsidence to any definite volume of ore removed. After due consideration of all factors, there has been selected for discussion an area 1400 ft. long by 1200 ft. wide, which will be referred to as the Queen Hill block. It is bounded on all four sides by faults, which tend to limit subsidence disturbance to that caused by ore mined from within the block, and likewise to prevent subsidence cracks due to mining outside of the block from entering the area.

Jan 1, 1934

By Rodney Lawrence, Tim Hackett

INTRODUCTION The western USA contains significant coal reserves in seams 10 feet or more in thickness which lie too deep for surface mining. As part of a demonstration of the use of two legged shield supports in the United States, Kaiser Steel Corporation, under contract to the US government used the longwall method to mine a ten foot thick coal seam in New Mexico. Caliper type shield supports and double drum shearers were used in three longwall panels. During the test, technical, production, and safety data were collected to objectively evaluate the performance of the shields in mining the thick coal seam. This paper presents discussions of practical operating experience and summary statistics to aid the mining industry in planning longwall mines in similar conditions.

Jan 1, 1981

By W. H. Emerson

In the summer of 1894, a specimen of corundum from Acworth, Ga., which mas reputed to be of markedly inferior quality for the manufacture of corundum-wheels, was received by the Geological Survey of Georgia, with the request that it be analyzed, in order to ascertain whether this inferiority was associated with any peculiarity of composition. In its appearance there was nothing uncommon, though it seemed to be rather easily scratched with a knife, while, at the same time, it scratched agate easily. An analysis gave the following results: Alumina, 94.58; silica, 1.77; magnetic oxide of iron, 0.69; calcium oxide, 0.44; and water, 2.51 per cent. There is nothing unusual in this composition, except the amount of water, which is usually, in southern corundum at least, from 0.3 to about 1 per cent. Concerning this corundum, the Mineral Resources of the United States, 1893, pp. 676 and 677, Bays: "It is supposed that this variety of corundum contains a little water, as it is

Jan 1, 1900

By Linda P. Bono

Some mining engineers are pursuing their destinies in places far removed from what they might have envisioned back in the classroom. The path they follow to work each day may wind down Wall St., Park Avenue, LaSalle, or Market; their offices are in modern skyscrapers which offer all the amenities-air conditioning, thick carpets, attractive decor, and executive lunchrooms, to name just a few. The papers on their desks are a combination of financial figures and feasibility studies. But even in these surroundings there are indications of what these men do and where their professional interests lie-minerals are everywhere, from vast collections to one symbolic paperweight. . . there are sometimes specks of dirt under fingernails from a recent on-site inspection, belying the traditional banker's garb. . . and there are the wistful comments about finding it necessary to get out in the field.

Jan 9, 1978

By Willis Carrier

PARTICULAR interest in the ventilation of deep mines, especially those in South Africa, has been created by a very complete system of cooling of the world's deepest mine, the Turf shaft of the Robinson Deep ill Johan-nesburg. The writer visited this mine in January, 1936, and again in 1937. During these visits and subsequently, a considerable part of the theory involved in rock cooling in tunnels was developed and this has been checked by practical observations of the rates of rock cooling made within mines. The purpose of this paper is to set forth the method of attack of the mine problem. The cooling or air conditioning of mines by the use of mechanical refrigeration was not wholly new 1-5 when the Robinson Deep installation was made. The first mine to be cooled was the Morro Velho mine, in Brazil. The first installation in this mine was put into operation in December 1920, the refrigerant being ammonia.5 The surface cooling gave considerable benefit but it proved inadequate as greater depths were reached, therefore in 1929 two centrifugal refrigerating machines were installed at the 6000-ft. level. These were used in conjunction with the surface cooling system to furnish additional cooling at the lower levels. This later installation used Carrene as a refrigerant. This nonflammable refrigerant, having a boiling point of 104° at sea level, insured reasonable safety, since all parts of the machine operated At subatmospheric pressure. Very favorable reports have been received on the effectiveness of this installation, but no recent data are available. An interesting feature of that plant is that the condenser water is cooled by spraying it into the exhaust air from the mine at the base of the ventilation upcast, which permits the water to be cooled within a few degrees of the wet-bulb temperature of the air entering the upcast, and the only water that has to be supplied for condensing purposes is that lost by evaporation, or about 0.5 per cent of the volume circulating through the condenser. This plan affords an ideal means for the removal of the heat load from the refrigerating system, and of course also from the mine itself. In many mines, however, especially those in South Africa, there is no

Jan 1, 1938

By Dan Chun

Abstract-To facilitate the efficient handling of large volumes of information generated by logging exploration drill cores, a computer data bank system (GEOBANK) has been developed to store and retrieve the normally difficult to handle geologic information as well as the physical testing data. The data bank systematized the logging of drill core by first developing a standard list of adjectives or descriptors (with the corresponding assigned numerical codes) for the various stratigraphic categories that are also assigned numerical codes. The system uses the "floating-slot" concept and handles the data bank with one single input form.

Jan 9, 1978

The Tooele plant is situated at the mouth of Pine Canyon, five miles northeast of Tooele City. The site forms one terminus of the Tooele Valley railway, which runs southwest through Tooele City to Warner, a station on the Los Angeles and Salt Lake Railroad, thirty-five miles southwest of Salt Lake City. An aerial tramway, four miles in length, crosses the Oquirrh mountains and connects the plant with the Bingham district. The plant consists of a lead-zinc flotation concentrator, lead smelter and copper smelter.

Jan 1, 1925