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By Eugene Rabinowitch

The Chemistry of Uranium. By Joseph J, Katz and Eugene Rabinowitch. Published by McGraw-Hill, Book Co. of New York, 1951. 609 P. $7.15.--This volume is the first half of a comprehensive treatise on uranium chemistry. It contains a detailed summary of the properties of the element uranium, including its occurrence in nature, extraction from ores, and the physical and chemical properties of uranium metal. Binary compounds of uranium with hydrogen, oxygen, nitrogen, and the halogens, among others, receive extensive discussion, including methods of preparation and physical properties such as crystal structure, thermochemistry and volatility. A very large proportion of the material in this volume is based on the results obtained in the laboratories of the Atomic Energy Project, much of it not published before. In addition, all of the open literature has been critically surveyed and integrated with the results of the Atomic Energy Project studies. Principles of Geology. By James Gilluly, Aaron C. Waters and A. O. Woodford. Published by W. H. Freeman & Co., San Francisco, Calif., 1951.640 P. $5.75.-This book attempts to summarize some of the knowledge that geologists have won from the study of the earth. It covers such subjects as movements of the earth's crust, records of earth movement, stream erosion and deposition, deserts, the ocean and mineral resources of the earth. The authors have concentrated on the analysis of processes that are at work upon and within the earth. rather than to present a catalog of descriptive facts and terms. The development of stratigraphy is also outlined. Pictures, maps, and tables are used to illustrate the major points of each chapter.

Jan 9, 1951

By F. H. Parsons

At the time that converting or bessemerizing of copper matte first began to be practiced by the smelters in the copper industry, converters were lined by tamping the flux, usually siliceous ore, into the steel shell. About 1910, practice in several plants had demonstrated the superiority of a basic lining of nonfluxing material and the addition of flux through the mouth of the converter. Adoption of the new method by nearly the entire industry followed. Basic linings are made of refractory "magnesite" brick manufactured, usually, from imported dead-burned magnesite. Either the linings are laid up in the shell dry or the bricks arc bonded with some mortar dictated by the preference of the user. Foreign, usually Austrian, magnesite was the main source of supply for these bricks, although there were domestic sources, but it was not until costs began to rise a few years after the first use of basic linings that these domestic sources were investigated to any great extent. Exigencies created by the World War necessitated the use of the domestic magnesite during this period, but the brick manufacturers reverted to the use of the foreign product after the war, because the impurities therein made a brick with a better fire bond than the purer domestic magnesite. Therefore the domestic product became available at a favorable price for any metallurgical process that permitted its use. The United Verde converting practice using magnesite linings has resulted in a high consumption of magnesite because of the zinc content of the low-grade matte. It was realized that the cost of these linings represented a large portion of the cost of converting, and an investigation was started to find means by which these costs might be lowered. In the neighboring state of California rather pure magnesite was being mined and treated for the market. The product, however, was purposely made to simulate the dead-burned grain magnesite product from foreign sources by the addition of fluxing impurities before burning. It was learned that this dead-burned California grain magnesite could be laid down in Arizona at a price one-third that of the same tonnage of "magnesite" brick manufactured from foreign magnesite. It was also learned that at least one California producer was ready to supply the market with periclase, which was a grain magnesia produced by burning the

Jan 1, 1934

By T. Toda, R. Maddin

A study has been made of the microstructure and mechanical properties of splat-cooled aluminum-base gold alloys with gold concentration from 0.25 to 5.0 wt pct. These alloys have been quenched from the liquid state by a torsion-catapult technique, which has made it possible to pepare specimens suitable for mechanical property measwements. From the electron micrographs it has been shown that the solid solubility of gold in aluminum can be extended to 2.5 wt pct (0.35 at. pct) by splat-cooling, while the maximum equilibrium solubility is known to be less than 0.3 wt pct (0.04 at. pct). The very fine grain size (several tenths of a micron), the extended solid solubility, and the fine dispersion of a second phase (AuAl2) contribute concurrently to a substantial strengthening effect. In Al-5 wt pct Au splat-cooled specimens of less than 50 thickness, the yield strength is 17 kg per sq mm or 6 times as large as the strength of bulk specimens. For the Al-1.0 to 2.5 wt pct Au solid solution obtained by splat-cooling, the yield strength reaches 7.5 kg per sq mm after an aging treatment (for 10 hr at 200°C), while it is 3.7 kg per sq mm for the corresponding bulk specimens. A great deal of research has been done in recent years on the structure and the properties of metals and alloys rapidly quenched from the liquid state.' The term "splat-cool" has been used with the meaning of a rapid quenching from the liquid state., The splat-cooling techniques have produced large numbers of new structures, which are expressed in terms of metastable phases,3 concentrated solid solutions,4 amorphous phases,5'6 new phases,7 and so forth. Nearly all previous studies have concentrated on the physical properties; i.e., crystallography, structure, electrical resistivity, magnetism, and so forth, of the splat-cooled metals and alloys. The mechanical strength of splat-cooled metals and alloys has hardly been investigated except for some recent work by MOSS' on A1-V alloys. The principle common to all experimental techniques developed to obtain very rapid quenching rates is based on the heat transfer by conduction. Liquid must be in good thermal contact with a substrate of high heat conductivity. Both of the published devices known as the "gun" and the "piston and anvil" techniques suffer from certain shortcomings. For example, the specimen obtained by the gun technique is very small and flaky, and hence inadequate for mechanical properties measurements. On the other hand if the material is forced to yield a continuous speci- men by the piston and anvil technique, it is probable that some plastic deformation occurs during the quench. A novel method for rapid quenching of a liquid metal or alloy, the "torsion-catapult", has been devised by Roberge and Herman9 at the University of Pennsylvania. In the apparatus the melt is thrown out of a curved furnace by a catapult and impinges on a copper substrate. The apparatus has the advantage of producing a continuous foil which is relatively large in size and of a quality suitable for the measurements of mechanical properties. The quenching rate is estimated to be of the order of l05 to l06 ºC per sec, (comparable to rates achieved by the piston and anvil technique). In selecting an alloy to be studied we were made aware of the fact that gold was believed to be "insoluble" in in and consequently age hardening in the A1-Au system appeared to be interesting. Quite recently Heirnendahl13-15 revealed that the solid solubility, as determined by transmission electron microscopy, was 0.3 wt pct Au at 640°C and 0.25 wt pct Au at 600°C, decreasing with decreasing temperature. In an A1-0.2 pct Au alloy after quenching from a solution treating temperature of 600°C the yield stress was 2 kg per sq mm, and it increased up to 6 kg per sq mm after aging for 1 to 10 hr at 200°C. The precipitation occurred in the form of platelike particles mainly on (100) matrix planes. The intermediate phase n', the equilibrium phase n (AuAl2), and lattice relationships between both precipitates and the matrix were also investigated by electron microscopy. One of the purposes of the present research is to determine whether or not the solid solubility in this system, in which gold has a very small solubility in

Jan 1, 1970

By Anthony Folger, Charles E. Straub

Kansas produced 38,150,878 bbl. of oil in 1928 and 40,658,170 bbl. in 1929, thereby retaining its rank as fourth among the oil-producing states of this country. Production for 1928 was less than that for 1927 (Fig. 1) because of a decrease in the number of completions, but the average initial production per well (Table 1) shows an increase due to the completion of wells of comparatively large capacity in the new Valley Center pool of Sedgwick County. This increase in average initial production per well was further augmented during 1929 by completion of additional wells of comparatively large capacity in the Valley Center pool and the new Greenwich pool of Sedgwick County. Production for 1929 exceeded that for 1928 and threatened to exceed all records except that of 1918, however, production was prorated in

Jan 1, 1930

By E. J. Schmitz

The mining of the Triassic coals of the Richmond basin has been carried on, upon a larger or smaller scale, for more than a hundred years. Notwithstanding the close proximity of the field to a large city and to tide-water, assuring both a good home-market and cheap transportation, and notwithstanding the good quality of the fuel and the great thickness of the principal bed developed and worked, operations in this field have been conducted with but limited success. The basin comprises about 189 square miles, being from 24 to 31½ miles in length by 5½ to 10 miles in width. Not much more than one-tenth of this area has been explored in a practical and efficient way, and not over 1200 acres have been worked near and along the outcrops or edges of the coal. The practical mistakes which have been committed undoubtedly in a number of the mine-operations of the past, do not alone explain the unsatisfactory success and progress of the industry of this field ; nor can it be ascribed to the lack of local enterprise. Other reasons have to be looked for; and, as a principal one, the irregularity and the faulted conditions of the coal deposits must be pointed out. As a second cause, the indifference of the State government, which leaves the development of the field entirely to private enterprise, instead of aiding it by systematical and practical geological investigations, deserves to be mentioned. Nothing has been done by the State since the excellent work done more than fifty years ago by Professor Rogers, who could not at that time enter into the details of such

Jan 1, 1895

By F. C. Bond

SIZE of grinding media is one of the principal factors affecting efficiency and capacity of tumbling-type grinding mills. It is best determined for any particular installation by lengthy plant tests with carefully kept records. However, a method of calculating the proper sizes, based on correct theoretical principles and tested by experience, can be very helpful, both for new installations and for guiding existing operations. As a general principle, the proper size of the make-up grinding balls added to an operating mill is the size that will just break the largest feed particles. If the balls are too large the number of breaking contacts will be reduced and grinding capacity will suffer. Moreover, the amount of extreme fines produced by each contact will be increased, and size distribution of the ground product may be adversely affected. If the balls added are too small, grinding efficiency is decreased by wasted contacts that are too weak to break the particles nipped; these largest particles are gradually worn down in the mill by the progressive breakage of corners and edges. Ball rationing is the regular addition of make-up balls of more than one size. The largest balls added are aimed at the largest and hardest particles. However, the contacts are governed entirely by chance, and the probability of inefficient contacts of large balls with small particles, and of small balls with large particles, is as great as the desired contact of large balls with large particles. Ball rationing should be considered an adjunct or secondary modification of the principle of selecting the make-up ball size to break the largest particle present. Empirical Equation In 19521,2 the author presented the following emerical equation for the make-up ball size: B - ball, rod, or pebble diameter in inches. F = size in microns 80 pct of new feed passes. Wi - work index at the feed size F. Cs - percentage of mill critical speed. S — specific gravity of material being ground. D == mill diameter in feet inside liners. K - 200 for balls, 300 for rods, 100 for silica pebbles. Eq. 1 was derived by selecting the factors that apparently should influence make-up ball size selection and by considering plant experience with each factor. Even though Eq. 1 is completely empirical, it has been generally successful in selecting the proper size of make-up balls for specific operations. But an equation based on theoretical considerations should be used with more confidence and have wider application. The theoretical influence of each of the governing factors listed under Eq. 1 was accordingly considered in detail, as described below, and a theoretical equation for make-up ball sizes was derived. Derivation of Theoretical Equation Ball Size and Feed Size: The basis of this analysis is that the largest ball in a mill should be just sufficient to break the largest feed particle into several pieces, excluding occasional pieces of tramp oversize. In this article the size F which 80 pct passes is considered the criterion of the effective maximum particle feed size. The smallest dimension of the largest particles present controls their breaking strength. This dimension is approximately equal to F. As a starting point for the analysis it is assumed that a 1-in. steel ball will effectively grind material with 80 pct passing 1 mm, or with F- 1000µ or about 16 mesh. The breaking force exerted by a ball varies with its weight, or as the cube of its diameter R. The force in pounds per square inch required to break a particle varies as its cross-sectional area, or as its diameter squared. It follows that when a 1-in. ball breaks a 1-mm particle, a 2-in. ball will break a 4-mm particle, and a 3-in. ball a 9-mm particle. This is in accordance with practical experience, as well as being theoretically correct. Confirmation of this reasoning is supplied by the Third Theory of Comminution," which states that the work necessary to break a particle of diameter F varies as F. Since work equals force times distance, and the distance of deformation before breaka4e varies as F it follolvs that the breaking force should vary as F½ These relationships are expressed in Table I, with a 1-in. ball representing one unit of force and breaking a 1-mm particle. This establishes theoretically the general rule used in Eq. 1 that the ball size should vary as the square root of the particle size to be broken. Ball Size and Work Index: The work input W required per ton" varies as the work index Wi, and the

Jan 1, 1959

By P. A. Witherspoon, Y. W. Tsang

Normal stress-fracture closure variation and stress-dependent fluid flow rates were calculated for rock fractures of different aperture distributions. This study shows that both the mechanical and hydraulic properties of the fracture are controlled by the large-scale roughness of the rock fracture. Hence the typical large-scale undulation wavelength of the rock joint dictates the suitable sample size to be used in normal stress and fluid-flow measurements, if results due purely to sample size are to be avoided.

Jan 1, 1982

By E. C. Weichel, J. H. Carpenter, J. C. Detweiler, J. L. Gillson, J. P. Wood

ILMENITE, rutile, zircon, and other heavy minerals are found in small amounts in the sand that covers most of Florida. Small enriched deposits occur on the beaches and coastal dunes. One beach deposit east of Jacksonville was worked during World War I and another near Vero Beach during World War 11. Both ceased operation when the resumption of imports reduced prices.

Jan 8, 1953

By Charles A. Ashburner

One of the most unexpected and unusual mining disasters which have ever been recorded in the Pennsylvania anthracite region, or, in fact, in any coal-mining district, occurred in the northern anthracite field near Nanticoke, Luzerne County, about 10 o'clock on Friday morning, December 18, 1885. This accident caused the death of twenty-six persons, a larger number than had perished in any catastrophe in the anthracite mines since that which occurred at the West Pittston mine in 1871. The Nanticoke disaster resulted from a " cave-in " in the roof of a portion of the Ross coal-bed workings from Slope No. 1 of the Susquehanna Coal Company, by which a large body of quicksand, gravel, rock- and coal-boulders, and water broke through the roof of the mine, filling up a large extent of the workings, until not only the water, but also the rock material, flowed out at the mouth of the mine. Immediately over the point where the "cave-in" occurred, was located a large pile of refuse coal and slate, taken from Slope No. a. After the occurrence of the accident, I made a very careful examination of the ground in the vicinity of the "cave-in," in conjunction with a study of all the maps, sections and other mining data in the possession of the Susquehanna Coal Company, which were placed at my disposal by Manager Irving A. Stearns. The accompanying map, Fig. 1, was compiled from the Company's map, and from special maps made by the Geological Survey under the direction of my assistant, Mr. Frank A. Hill. This map shows the relative position of the mine-workings, the surface features and the "cave-in," and other points to be specially referred to hereafter. My examination was made particularly with the view of determining the indirect cause of the " cave-in," and the local geologi-

Jan 1, 1887

By R. W. Raymond

AT the Liége meeting of the Iron and Steel Institute of Great Britain, in August, 1873, and also at the Vienna Convention of Mining and Metallurgical Engineers, at the end of August, and at the convention of German geologists, at Wiesbaden, in the first part * Extract from the President's address.

Jan 1, 1874

By C. L. T. Edwards

IN the preparation of this paper, the author has drawn upon experience with the operation of a blast furnace on 100 per cent scrap burden, which he believes was the first operation of its kind in the country. It is the writer's further belief that much of the controversy attending discussions of this subject may be obviated by the presentation of a group of facts that should be known by blast-furnace men and open-hearth men alike. The following question has been raised: "Is iron made with scrap in the blast-furnace charge as good iron as that made without scrap-made with straight ore?" This question may be answered in part by the fact that the open-hearth men of the Bethlehem Steel Co. who used pig iron from a blast furnace that operated on 100 per cent steel scrap found no difficulty in the processing of such iron other than that to be expected from the known analysis. OPERATION WITH 100 PER CENT STEEL SCRAP During World War I, there was accumulated at the Bethlehem plant of the Bethlehem Steel Co. some 22,000 tons of alloy-steel turnings and borings. This scrap material was highly corroded and, like alloy-steel turnings of that period, were as a drug on the market. There was also a small tonnage of scale, rich in valuable metal. If this scrap could be worked up into pig iron, it would become readily available for use in the open hearth. In order that the pig iron should contain the maximum percentage of nickel, permission was obtained from the Management to try a 100 per cent scrap operation. Accordingly, a blast furnace was swung over on July 25, 1919; the percentage of steel scrap increased more or less irregularly until July 29, when l00 per cent steel turnings were charged. Except for occasional charges of nickel scale as an "extra," the operation continued uninterruptedly until Aug. 27, when the furnace was swung back to the regular ore-ore and scrap-operation. During this period, the furnace production ranged from a low of 568.60 tons to a maximum of 766.07 tons. At the swing-over from i00 per cent scrap to the regular burden, the charge hung up rather severely. While the latter condition has nothing to do with the object of the paper, a brief discussion of the experience will appear later because of the interest it may have for blast-furnace operators. In Table I, results of the first and last days of the ore charge are shown, July 25 and July 29. The last column shows a summary for the period of maximum scrap charge. The furnace employed had a hearth 16 ft. 0 in., bosh 22 ft. 0 in., stock line 15 ft. 6 in., bell 11 ft. 6 in., batter ¾ in., bosh angle 73° 27' and a Baker-Neuman top, with gas-engine blowers. Analyses Range of Material Charged Gravel: Si02, 64.40 to .7o.54 per cent; A1203,3-52 to 4.58; CaO, 7.27 to 9.49; M90, 5.09 to 6.14; FeO, 1.67 to 3.34, ignition loss, 9.50 to 13.83.

Jan 1, 1941

By R. B. Snow

Inclusions revealed by the ultrasonic inspection of forgings, slabs, and blooms cause costly diversion or rejection of the product. Most of those inclusions are so large that they should have floated out of the liquid steel soon after teeming and been removed when the top discard was taken from the ingot. To investigate the source of the relatively large inclusions found in the body portions of ingots, a large forging ingot was examined from top to bottom along the centerline, and special studies were made of open-hearth and BOP ingots. A study of the inclusion distribution in the as-cast 134-inch-diameter forging ingot revealed only a few inclusions larger than 15 microns in the body of the ingot. Larger inclusions may be formed by entrapment of rising inclusions in large falling dendrites rather than by the precipitation and agglomeration of oxides from the liquid steel-dendrite interface during solidification. The original top crust of the ingot investigated broke away from the hot top and sank to the bottom during the solidification process, as indicated by entrapped inclusions found in a band 12-1/2 inches above the stool. A change in the hot-top insulation practice on subsequent ingots to insure the integrity of the top crust has produced clean ingots. A ZrO2 tracer was added to the hot-topping compound and to the top surface of open-hearth and BOP ingots after teeming. The tracer was found in large inclusions in slabs and blooms from these ingots. Large inclusions in the body of ingots appear to be related to disturbances of the natural freezing and shrinkage processes in the hot- top region of the ingot.

Jan 1, 1972

By Hugh H. Waesche

AN adequate supply of crystalline quartz of piezoelectric grade and size continues to be of fundamental importance to the U. S. Army Signal Corps. Current electronic development programs of the Armed Forces depend upon the use of quartz crystals in all applications where accurate frequency control is necessary. In peacetime this' may not seem too significant, but in planning for possible national defense emergencies, an adequate supply of piezoelectric grade raw quartz is just as important as an adequate supply of some of the more publicized strategic minerals. Prior to World War II, the Signal Corps accepted the use of quartz crystals for frequency-control purposes in essentially all electronic communication, navigation, and radar equipment operating in the audio and radio frequency spectra up to and including the ultra-high-frequency channels. Over 4,000; 000 lb of radio-grade quartz was required between January 1942 and V.J Day. Use of such a large quantity of crystal units was necessary in order to establish and maintain, with a minimum of effort, clear and definite frequency channels of greater stability than could be obtained by any other means. This policy was justified on the basis that quartz was without a peer as a frequency-controlling element. Military and civilian uses of quartz in the field, as well as laboratory studies, have confirmed this premise and no means of highly stable frequency control, superior to quartz, has been discovered. It appears certain that the demand for use of quartz for frequency control purposes will continue at a high level.

Jan 1, 1948

By E. S. Wheeler

The conversion plant of the Clirnax Molybdenum Co. is at Langeloth, Washington . County, Pennsylvania, approximately 30 miles west of Pittsburgh. The molybdenite concentrates converted originate in the company's mine and mill at Climax, Co1orado.l They are delivered to the Langeloth plant by rail packed in 175-lb. bags. War production of concentrates at the mill started in 1918 but it was not until about 1920 that any peacetime uses of the metal began to develop. Early conversion of Climax concentrates was mainly into ferromolybden um, produced for the company by a custom smelter. Chronology of Plant Development In 1924 a small hand-rabbled roaster was built at Langeloth, the roasted concentrates being used principally in the production of calcium molybdate by a conventional wet method. The size of the original plant is indicated by the fact that during 1925 about 400,000 lb. of molybdenum was converted. In 1925-1926, a multiple-hearth roaster was erected (12-ft. 0.d. 8-hearth Nichols-Herreshoff); also the first experimental ferromolybdenum was produced at Langeloth. At this time sa.1cium molybdate was produced in the furnace by a process to be described. With continued interest in and increased use of molybdenum, it was necessary to arrange for enlargement of plant facilities. Although only one 12-hearth 16-ft. 0.d. roaster was erected at this time (1929), plans were prepared for a duplicate unit, which was completed in 193 5, giving a potential roasting capacity of some 16,000 lb. Mo per day under the then prevailing roasting practice. Shortly after this second unit had gone into production, increased demand overtaxed the plant, and as a result two 12-hearth 18-ft. 0.d. roasters were placed in operation in 1936 and 1937. This gave the plant a potential roasting capacity of 40,000 to 50,000 lb. Mo per day, and necessitated major changes in auxiliary handling equipment as well as improved facilities for producing ferromolybdenum. Following the introduction (1938) of a new product—molybdic oxide briquets— and because of the constantly growing demand for molybdenum, plant facilities were again enlarged in 1940 and 1941 by the erection of two 16-hearth 18-ft. 0.d. roasters. These two new units, along with a change in roasting practice, brought the potential roasting capacity to 150,ooo to 190,000 Ib. Mo per day, based on about 15 lb. per sq. ft. of hearth area. Other departments of the plant have necessarily grown, until today the plant is modern throughout and completely mechanized, with equipment geared to produce marketable products as follows:

Jan 1, 1944

By B. J. Kochanowsky

The idea of a portable crusher is not new. Many such crushers are available but they are small and designed for construction work. For many years the author has suggested, both in this country and in Europe, the building of larger portable crushers intended expressly for use in quarries or open pits. Although not applicable under all conditions, there are mining operations where a mobile crusher arrangement could be more profitable than the facilities now used. The primary use of a portable crusher, i.e., a crusher mounted on crawlers or tires, in the rock and mining industries is to reduce costs by permitting the substitution of conveyor belt haulage for truck or track haulage. The usual sequence of operations in surface mining is drilling, blasting, loading, haulage, and crushing. Haulage is normally accomplished by truck or track-mounted cars, the latter method being used for the longer distances. However, by using a portable crusher in the pit, the sequence of operations would be changed so that the crushing stage would occur before haulage (Fig. 1). Such a sequence would permit the use of conveyors to replace the more expensive truck or track haulage methods. Since most quarry and open pit operations normally require a crushing stage, the only additional costs incurred will be due to the investment required to purchase or construct a mobile arrangement for a crusher. But this factor has to be weighed against the advantages to be gained by conveyor haulage. As shown in Fig. 2, transportation of material by belt conveyor over short distances is less expensive than by truck. The inclination of the belt has no effect on belt speed; consequently, the hourly tonnage moved remains the same. Conversely, the output rate of trucks as expressed in tons or ton-miles per shift decreases proportionally to the haulage speed, which is considerably slowed by the steepness of the road (Fig. 3, left). Although maximum possible grades and maximum economic grades of haulage are greater for a belt than for a truck (over the same total lift), the longer haulage distances favor the use of trucks. Although power consumption for hauling on a grade increases for both conveyances, the rate of power consumption increases faster for trucks than for conveyor belts (Fig. 3, right). Since the output rate and related fixed costs are affected by the travel speed, total haulage costs with trucks would increase with the grade more rapidly than the similar costs of conveyor belts (Fig. 4). Travel distance, road grade, speed, size and number of pieces of equipment, efficiency of operation, and many other factors affect such haulage costs. In general terms it can be said that the shorter the distance, the steeper the grade, and the greater the output, the more advantageous the belt becomes in comparison to truck or track haulage. In addition to potential cost savings in haulage procedures, a portable crusher would allow better utilization and performance of shovels. Loading operations would not be interrupted as often by the necessity of waiting for cars or trucks. Unfortunately, the application of belts in open pits for haulage from bench sites is generally not practical under existing conditions because a belt fed directly by a mechanical shovel can be torn, damaged, or worn out quickly by the large rock fragments falling on it during loading. However, by using a mobile crusher this situation can be avoided. As shown in Fig. 1 (b), the shovel feeds rock into the crusher located behind it. The crushed material is initially transported by an extensible and/or movable belt, thence by a longer stationary conveyor to the plant where the material is subjected to further treatment by secondary crushing, screening, etc. The first-mentioned conveyor, needed to bridge the distance between the shovel and the stationary conveyor, is necessarily variable in length owing to the continuous movement of the shovel and the desire to keep the stationary belt at a safe distance from the bench during blasting operations. The remarkable part of mobile crusher operations is the extra-ordinarily high output per man-shift, the low maintenance and power requirements for haulage, and the increased output of the loading shovel. A cement quarry which has been using a portable crusher and conveyor since 1956 requires only three men to operate the shovel and crusher and to transport the crushed rock by belt from the quarry face to the screening plant. If truck haulage

Jan 1, 1961

One of the papers under this head covers the sampling and estimating of iron ores of the Lake Superior region as a whole while the other four papers cover the Gogebic, Marquette, and Menominee Ranges of Michigan and the Cuyuma Range of Minnesota.

Jan 1, 1925

By K. B. McQuillin

The increasing volume of data produced by the Anglo American Corporation of South Africa, Limited, during the course of coal exploration prompted the establishment of a computer based processing system for coal data. This system, known as the Coal Mine Evaluation System, consists of geological, mine planning and financial evaluation sub-systems. This paper describes the use of the geological sub-system in assessing the surface mine potential of a portion of the Witbank Coalfield, South Africa, in which more than one thousand core boreholes have been drilled to date.

Jan 1, 1977

By Wilbur Nelson

THE year 1924 showed even less drilling in Kentucky and Tennessee than the preceding year. The operators in this section, however, are anxious to resume development work and are only waiting for a higher market for crude oil before starting such a program. In Kentucky, during 1924, 1605 wells were completed, the combined initial daily production of these wells being 25,599 bbl. Of the total number of wells drilled 1176 were oil wells, 61 gas wells, and 368 were dry holes. The average initial production of the oil wells was 21.8 bbl. per well. In Tennessee, 28 wells having an initial daily production of 1199 bbl. were completed during 1924. Of these, 3 were oil wells, 6 were gas wells, and 19 were dry holes. The initial production of the three oil wells was 400 bbl. Of the total number of wells drilled in Kentucky in 1924, 73 per cent. came in as producing wells; and of those drilled in Tennessee, 10 ½ per cent. came in as producing wells and 20 per cent. as gas wells. In Tennessee, during 1924, wildcatting dropped off considerably and only about one-half as many wells were drilled as in the preceding year; in Kentucky, on the other hand, there was only a slight curtailment in drilling, there being 1943 wells drilled in 1923 as against 1605 in 1924.

Jan 3, 1925

By Cleveland Meeting

Jan 1, 1945

By J. S. Levine

A series of four displacement experiments has been run in a large alundum core. Flow potential distribution in each liquid phase was measured continuously through oil-wet and water-wet capillary barriers, using a null-point pressure balance, and saturated distribution was determined from electrical resistivity measurements. A method has been devised for calculating the effective (and relative) permeability to each phase at any time and at any point in the flow system for the transient displacement process. Relative permeability points have been calculated and average curves drawn for oil displacing water and water displacing oil. The values of dynamic capillary pressure have also been determined. Data have been substituted into the generalized Buckley-Leverett fractional flow equation and a comparison between actual and theoretical breakthrough recoveries is shown. Displacement experiments have been performed in the laboratory for many years in an effort to determine the effect of a number of parameters on multiphase flow in porous materials. In the literature of petroleum technology, Wyckoff and Botset' were among the first to develop the relative permeability concept and to show experimentally that oil, gas, and water can flow simultaneously through a porous medium in a precise way governed by certain fluid properties, by the relative saturation of each fluid, and by characteristics of the medium itself. Somewhat later, Leverett2 and Hassler, Brunner, and Deahl11 developed some theoretical concepts of capillary behavior in porous materials. The work of Buckley and Leverett3 presented a theoretical analysis of immiscible liquid displacement based on the equation of continuity and the Darcy equation for viscous flow of each fluid. Subsequent investigators attempting to verify the Buckley-Leverett equation have, in general, deliberately neglected the capillary pressure term or have performed experiments in such a manner that it could be neglected. Terwilliger, et al4 in 1951, obtained agreement between theory and experiment for the constant rate displacement of water by gas in a long vertical sand column. Rapoport and Leas5 have investigated the

Jan 1, 1955