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By Lloyd Williams

1. This investigation shows that the Sewanee bed is the only coal bed being mined in Sequatchie County. Other coal beds are present, but as they do not contain known recoverable reserves, estimates of reserves were made only for the Sewanee bed. 2. Known measured and indicated reserves in the Sewanee bed, based on a minimum thickness of 14 inches and on 1,800 short tons per acre-foot of coal in place, are estimated to be 76 million short tons as of January 1, 1953. Of this total 62 million short tons represent coal 28 inches and more thick. Areas in the bed were omitted from the estimate because available data relative to the bed characteristics were insufficient to permit making an estimate conforming with the definitions of measured and indicated coal adopted for this study. Should future drilling or development prove reserves in these areas, such reserves should be added to the total estimated reserves determined by this investigation.

Jan 1, 1955

By R. C. Buehl

Pilot-plant tests have been in progress at the Bureau of Mines, Pittsburgh, Pa., on the recovery of manganese from open-hearth slag.3/4/5/ In the process, 1 the slag is first reduced in a blast furnace to make a high-phosphorus spiegeleisen containing 12 to 22 percent manganese. A slag containing 55 to 60 percent manganese is made from this spiegel by oxidation in a basic-lined converter. The metal from this blowing operation contains about 3 percent phosphorus, 1 percent manganese, and less than 1 percent carbon. Economic efficiency in the manganese-recovery process requires that this metal be converted into steel. Because of its high phosphorus content the metal is suitable for blowing in a basic-lined converter. In most of the tests described in this report, metal of this composition was used as the starting material in a 1-ton, side-blown, basic-lined converter. Although basic-lined converters are not used for steel production in the United states, in Europe a considerable portion of the steel is made in them. Open hearths largely have replaced converters for steelmaking in the United States; mainly because open-hearth steels have better physical properties than converter steels because of lower nitrogen and phosphorus contents. However, by improving converter design and operating technique during the past 15 years, metallurgists have found that low-nitrogen steels with physical properties equal to open-hearth steels can be produced in converters.6/ Tests at Jones & Laughlin steel Corp. with side-blown, acid-lined converters of 3 and 22 tons capacity showed that steel with a low nitrogen content comparable to that of open hearths operating on scrap charges could be made by surface blowing.7/

Jan 1, 1955

By R. H. Frazier, F. W. Boulger

THE performance of welded structures is closely associated with the ductile-to-brittle transition temperature of the steel from which they are made. A low transition temperature is desirable because it indicates that the steel is less likely to fail suddenly at low ambient temperatures. Structures such as bridges, ships, storage tanks, and pipelines are usually made from hot rolled semikilled steel. Changes in rolling practice or chemical composition appear to be the most practical methods for improving the toughness of such materials. This results from the fact that production is likely to be seriously curtailed if improvements were obtained by recourse to heat treatment or complete deoxidation. This paper discusses the effect of variations in carbon and manganese contents on the properties of semikilled steels. The transition temperature, a property to which considerable importance is attached, varies with specimen configuration, testing method, and criterion of performance. Both the Navy tear test' and the keyhole Charpy test were used in the investigation. According to the terms used by Vanderbeck and Gensamer,V he tear test was used to measure a fracture transition and the Charpy test to measure a ductility transition. In either case the specimens absorb considerably less breaking energy in tests below the transition temperature than in tests above the transition temperature. Decreasing the testing temperature of notched-bar specimens seems to be equivalent in its effect to increasing the severity of loading on fabricated structures. Therefore, structures built from steels exhibiting lower transition temperatures in laboratory tests are expected to be less susceptible to sudden brittle fractures in service. Consequently, changes in composition which lowered the transition temperature of the experimental steels were judged desirable. Materials and Methods The steels for this study were made in a laboratory induction furnace and rolled to 3/4 in. plate, using a finishing temperature of 1850°F. Precautions taken to insure reproducible melting and testing practices are discussed in detail elsewhere.' All tests were made on hot rolled steels. The analyses of the experimental steels are given in Table I. The list includes steels with manganese contents ranging from 0.21 to 1.46 pct, at each of five carbon levels. The phosphorus, silicon, sulphur, and nitrogen contents of the steels are reasonably constant. The tensile and notched-bar properties of the experimental steels are presented in Table 11. Both upper and lower yield points are listed, and all values are averages for duplicate specimens. Transition temperatures based on three criteria are given. The keyhole Charpy transition temperatures are the temperatures at which the energy-temperature curves based on averages of four specimens cross the 20 and 12 ft-lb levels. Both definitions of Charpy transition temperatures have been used by other investigators. Charpy tests were made at intervals of 10°F, using a pendulum with an available striking energy of 220 ft-lb and a velocity of 18.1 ft per sec. The specimens were oriented parallel to the rolling direction and notched normal to the surface of the plates. The tear test transition temperatures are based on the usual' criterion of the highest temperature at which at least one of four specimens exhibits brittle behavior. The dimensions of the tear test specimens are shown in Fig. 1. The specimen is loaded eccentrically in tension, with pin and shackle fixtures, through the large holes while submerged in a liquid bath at the proper temperature. A specimen developing a fracture area with less than 50 pct shear or ductile texture is classed as brittle. Tear tests were made at intervals of 10°F.

Jan 1, 1955

By L. D. Jaffe

THE data recently published1-" on heat treatment of titanium alloys have not been readily systematized into an overall picture. When the ß titanium phase transforms isothermally at low temperatures, the hardness follows a typical age-hardening pattern, going through a peak as schematized by the "hardness" vs "time of isothermal transformation" coordinates of Fig. 1. As the temperature is raised, the peak will decrease in height and occur at shorter times; slowing must however take place as the /3 transus is approached. The brittle microconstituent in the region of peak hardness has been termed /3'. Its structure is still undetermined but it is clearly a transition stage from soft /3 to the £ + a that is found at the right of the peak. On continuous cooling, high cooling rates correspond to short times on Fig. 1; abscissa labeling has been added accordingly. Extremely rapid cooling will retain soft /3; very slow cooling will precipitate soft a; intermediate rates give hard 8'. Increasing distance from the water-quenched end of a Jominy hardenability bar corresponds to slower cooling (right on Fig. I), but the range of cooling rates covered by a Jominy bar is only a small portion of that shown in the figure. The p-stabilizing alloying elements retard the transformation process, moving the material to the left on Fig. 1. A lean alloy is harder water-quenched (P') than air-cooled (/3 +a); a rich alloy is harder air-cooled (/3') than water-quenched (/3). Examples of the position of various compositions with respect to the curve, based on limited data in the literature, are given at the top of the figure. Oxygen and carbon accelerate the transformation process, moving the material to the right on Fig. 1. If an alloy quenched quickly enough to retain soft /3 is reheated below the £ transus (tempered), it ages, going through a hardness maximum as time increases (moving right on Fig. 1). Material which after cooling from the /3 range lies on or right of the hardness peak, will move still further right on tempering, and so soften. Since the rate of the tempering (aging) reaction increases with temperature, the figure also may be used to represent the variation in properties with tempering temperature, at short tempering times. The simple picture given here neglects martensite transformation, as well as differences between decomposition products of /3 formed by direct isothermal transformation and those formed by quenching and tempering. It may nevertheless be of use as a guide for practical heat treatment and for further research.

Jan 1, 1955

By W. A. Olds, H. F. Yancey, M. R. Greer

John Griffen (Consulting Engineer, Pittsburgh)— This paper presents much data offering tentative answers to many technical questions that have been raised by the employment of dense-medium processes and, particularly. by use of the drum-type vessel. It is to be hoped that the Bureau of Mines will publish complete detailed data for the 50 or more tests so that it may be studied and analyzed by everyone interested. Such study will take considerable time and effort. The writer is therefore obliged to confine his discussion to a few points he has noted as being open to question. The data in Table VI which shows the recovery efficiency for the 1 to ½-in. fraction as 100 pct raised a question concerning the accuracy of these calculated figures; washability data on the washed coal and refuse show that such a figure is impossible. When the ash content of the misplaced material in the washed coal is higher than that of the misplaced material in the refuse, recovery efficiency must be less than 100 pct. A etudy of the data In Tables 111, IV, V, and VI suggest that this error probably arose from slight errors in calculating the data in Table V and more probably from the small scale used in plotting the washability curves from the Table V data. Using the size consist data and the weight and ash percentages of individual gravity fractions of each size as given in Tables 111 and IV and the reported recovery of 80 pct of washed coal, a revised Table V was calculated, carrying weight and ash percentages to hundredths of a percent instead of rounding off at tenths. This washability data for each size of Belling-ham reconstituted feed was then plotted on large scale washability curves. From this data the new Table XIV was prepared, which agrees very closely with data shown in Table VI except for two important values, theoretical yields and recovery efficiencies. As will be noted the recovery efficiencies in Table XIV are significantly lower than those given in Table VI. A study was then made of the data showing distribution of the specific gravity fractions to washed coal and refuse. This study indicated that when this distribution was based on the fractions as tested at 0.1-sp gr intervals the number of points obtained were insufficient to establish the distribution curves with reasonable precision, which in turn would affect the accuracy of the values for the specific gravity of separation and the error areas. The data in Tables III and IV of the paper and curves prepared therefrom make it possible for the specific gravity fractions to be subdivided further. This was done using 0.05-sp gr fractions and the distribution data so obtained is given in Table XIV, with resulting values for the specific gravity of separation and the error areas. The writer believes that the derived values shown on Table XIV are more accurate than the correspond-

Jan 1, 1955

By S. H. Ash

THIS REPORT covers the third in a series of five projects into which the Conowingo tunnel system has been divided for geographical and practical reasons. It outlines the work constituting Project No.3 that consists principally of the following: (a) To sink the No. 15 shaft and extend the Alden, Sugar Notch, Baltimore No.2, and Laflin shafts from their present depths to tunnel level; (b) to drive a tunnel off these shafts to connect with the western end of the tunnel in Project No.1 (Lackawanna) and to a point 1 mile south of the No. 15 shaft; (c) to construct an emergency pumping plant at the No. 15 shaft, complete with 3 centrifugal pumps of 200-cubic-foot-per-second (c. f. s.) capacity each, and a discharge-water shaft and a discharge tunnel for conducting the water from the pumping plant into the Susquehanna River; and (d) to construct a lateral tunnel 1 mile long off the main tunnel in the vicinity of the No. 15 shaft, to be known as the Wyoming lateral. The project, as outlined, is a self-contained drainage unit that will permit unwatering inundated mines and safeguarding active mines in the Wyoming Basin of the Northern field and also will handle the water in the Lackawanna Basin, thus permitting the temporary pumping plant provided in Project No.1 to be discontinued. The pumps installed in the temporary pumping plant provided in Project No.1 can then be moved to the emergency pumping plants recommended in Projects Nos. 2 and 3, to be installed as the third or standby unit.

Jan 1, 1955

By D. R. Walker, M. C. Shaw

The laws of Kick and Rittinger are explained as functions of particle size with metal cutting theory. Comminution is shown to be basically the same process as metal grinding. The machine shop type of grinding operation is used to study mineral crushing. Evidence indicates that plastic flow occurs in comminution. THE laws of comminution of Kick and Rittinger have been debated for many years. Certain data obtained from ball mill and drop tests are found to be in approximate agreement with Rittinger's law while other data would seem to support Kick's law. In all these tests the energy required to produce a wide variety of particle sizes has been studied. Surface area has also been measured with considerable difficulty, at first because it was thought that the energy associated with crushing was that required to create new surface, later because of precedent and because a more fundamental variable was unknown. Tests made with the machine shop type of grinding process show that normally brittle materials that are usually crushed can behave like the more ductile metals. Brittleness is then a relative prop- erty largely dependent on specimen size and stress state. The machine grinding operation offers a convenient means for studying the comminution of materials in which the important size variable retains its identity. The energy required to grind a certain volume of material is shown to be strongly dependent on size. Kick's law is found to hold very well for particles below 1 micron in size, while Rittinger's law holds approximately in the comminution of larger particles. While a physical origin of Kick's law is established in terms of the theoretical strength of materials having a perfect lattice structure, there is found to be no physical basis for Rittinger's law. Because of its industrial importance, comminution has commanded considerable technical attention and has been the subject of continuous study for nearly a century. A number of empirical rules have been proposed to make it possible for grinding energies to be predicted. Of these so-called laws of comminution those bearing the names of Kick and Rittinger have been most widely applied and debated.

Jan 1, 1955

By Sherwin F. Kelly

"THE TITLE I have chosen is certainly not one of those self-explanatory ones. I shall endeavour to make its significance clear during my remarks but, in the hope of minimizing that inevitable exodus which follows a good lunch such as this, I am going to withhold the final key to its meaning until the very last sentence of this talk.Some explanation is due at the beginning, however, so for those of you unfamiliar with Spanish I will explain that casa means house, or home. The casa about which I am going to talk necessitates a wide .and generous interpretation of the word, because I am using it to indicate principally that portion of the North American continent which is the home, in the western hemisphere, of British and American ideals of justice and freedom, the dignity of the individual and his right to exercise private initiative .and engage in private enterprise. In a wider sense, I am including the entire western hemisphere, which may yet stand as the last bulwark of political and industrial democracy.As for the treasures of which I am about to speak, may I remind you that there are no• material treasures but those derived from the earth, and they are derived by one of two processes, -mining and farming. These two fundamental industries are the basis, and the sole material basis, of our welfare, our wealth, our very existence. From farming, whether it be farming of the land. or farming of the sea, we obtain the food to keep us alive, healthy and able to work."

Jan 1, 1955

By H. Reiss

The zone-melting process in which redistribution of solute in a solid bar is effected by the passage of a molten zone is considered mathematically. Simple approximate techniques are developed for computing the manner in which redistribution occurs as molten zones continue to pass. The introduction of the "zone-melting flux" provides valuable insight into the nature of the phenomenon, as well as a central mathematical theme in terms of which the process can be discussed. Curves are presented for typical and general cases. AN extremely efficient purification technique was described recently by Pfann.1 Known as zone refining, it consisted of the slow passage of a series of molten zones, produced by external heaters, through a long solid ingot of the substance to be purified. Equations were given for the maximum separation attainable, and a numerical computation method was described for obtaining the concentration in the ingot after a given number of zone passes. The present paper offers an interpretation of the physical nature of zone melting in terms of a transport process in which formal diffusive and convec-tive flows occur. This interpretation, in addition to providing insight into the physical nature of the process, also provides a basis for its mathematical treatment. From this basis, useful equations are developed for the solute concentration in the ingot as a function of the number of zone passes. These are applicable particularly where large numbers of passes are necessary and where numerical computational methods might be laborious. All considerations will deal with the limiting form of the process, described by Pfann,' in which: 1—Diffusion in the solid is considered negligible. 2—Mixing in the liquid is supposed, complete. 3—The distribution coefficient, k, measuring the ratio (at equilibrium) of solute concentration in the solid to that in the liquid is supposed, independent of composition. In general, this paper will be concerned with a bar of uniform (actually unit) cross section, ex- tending in the x-direction from x = 0 to x = L, where L may be infinite. Zone melting is a transport process, and as such must admit of interpretation in terms of a flux of transported quantity. However, the progress of operation is not measured in units of time but in terms of the number of completed zone passes. Consequently, the zone-melting flux will be a rate, per unit pass, rather than per unit time. The description of this flux and the examination of its properties is given below. Zone-Melting Flux Fig. 1 represents an intermediate stage of the nth pass. Concentration C is plotted against distance, x. The molten zone with its back at the point x and front at x + I moves in the direction indicated by the arrow. As it moves, it leaves behind the nth

Jan 1, 1955

By W. L. Crentz

The Black Mesa coal field is part of the Hopi and Navajo Indian Reservation in Apache, Navajo, and Coconino Counties, Ariz. The field is in an isolated area with the nearest railroad 70 to 90 miles to the south. Travel is restricted by roads that become impassable during inclement weather. The area is a mesa and rises from 500 to 2,000 feet above the surrounding territory. Water is very scarce, and the nearest stream of any consequence is the Colorado River about 90 miles northwest from the center of the field. Coal beds are found in the Mesaverde formation of the Cretaceous age, but they are uncorrelated and range from subbituminous to bituminous in rank. Huddle2/ estimates the reserves of subbituminous coal in Arizona at about 15 billion tons, most of which is in the Black Mesa field. The geology of the field has been reported by Campbell and Gregory.3/ Coal production is small, and the output is used almost exclusively for domestic fuel in the Indian schools, missions, and trading posts.

Jan 1, 1955

By S. H. Ash

THIS REPORT covers the second in a series of five projects into which the Conowingo tunnel system has been divided for geographical and practical reasons. It outlines the construction work contemplated-to sink two shafts on the line of the main tunnel that are designated the Nos. 9A and 11 shafts, to turn off from the No. 9A shaft the Lansford and Lykens laterals, and to drive portions of the main tunnel. An emergency pumping plant, consisting of 2 separate pumprooms, each containing 3 horizontal-type centrifugal pumps of 100,000-gallons-per-minute (g. p. m.) capacity, will be constructed at the No. 9A shaft. From the No. 11 shaft will be turned off the Trevorton lateral and portions of the main tunnel, which will connect with the section driven from the No. 9A shaft. Under the proposed project, the Lansford lateral will be driven for a distance of 2 miles, the Lykens lateral for a distance of 10.56 miles, and the Trevorton lateral for a distance of 20 miles. To drive the Lykens and Trevorton laterals for the distance indicated, several construction shafts and slopes must be sunk to permit efficient and safe construction.

Jan 1, 1955

By H. M. Fisher, R. E. Snow, S. C. Sun

RAPID colorimetric estimation of the amount of transparent minerals has been applied successfully to instream mill products at the cryolite flotation plant of the Pennsylvania Salt Mfg. Co., Na-trona, Pa. Apparatus used was a Bausch & Lomb monochromatic colorimeter. The optical assembly of the colorimeter, shown in Fig. 1, passes light from a 6-v, 32-cp, prefocus incandescent lamp through two condensing lenses, a heat-absorbing filter, an aperture-control diaphragm, an interference filter, and a sample container. The light finally strikes a barrier-layer photocell and the output current of the photocell is measured by a galvanometer. The engraved index line of the collective lens, located between the lamp and the galvanometer, is imaged upon the transmission scale of the colorimeter by the galvanometer mirror. Reflected light from the galvanometer mirror, the position of which is determined by the output current of the photocell, enables reading of percent transmission throughout the mineral sample being tested. The colorimeter, Fig. 1, was standardized as follows: 1—The colorimeter was' connected to a Sola constant voltage transformer, A, of 6-v output operated from a 110-v ac supply. 2—Turning the toggle switch, C, illuminated the transmission scale, B, and the control knob, D, was adjusted with a blank in the optical system until the hairline on the scale read zero. 3—A No. 430 interference filter, E, was inserted into the filter channel, F. 4—A test tube, G, containing 7 cc of distilled water was pressed into the sample space, H, and the fine adjustment knob,

Jan 1, 1955

By C. C. Popoff

The growth of the aluminum-smelting industry in the Pacific Northwest, resulting from national defense requirements, emphasized a need for local sources of alumina to supplement the alumina production shipped from the East. Because supplies of foreign bauxite were being interrupted and our domestic bauxite reserves were inadequate, it was necessary to expedite the investigation of other possible sources of alumina during World War II. A program of investigation of high-alumina clay deposits in the Northwest was carried out jointly by the Federal Geological Survey and the Federal Bureau of Mines. As part of this program the Cowlitz clay deposits near Castle Rock, Wash., were explored by the Bureau of Mines in 1942-43. These deposits are near railroads, tide-water, and Bonneville Power Administration lines and within 22 to 100 miles of 4 aluminum-reduction plants.

Jan 1, 1955

By J. D. Forrester, Erich Sarapuu, T. C. Cheasley

Electrocarbonization of coal will produce cheap fuel for electric power plants. The complete electrification of industry and domestic power consumption of gas generated as fuel in situ is a worthwhile goal for the coal industry. Production of synthetic oils, gasoline, and coal chemicals from coal gas will create additional new markets for the coal industry. RANKING high in the nation's wealth of natural resources, readily available liquid and gaseous hydrocarbons are nevertheless falling steadily in supply, while industrial requirements for hydrocarbon compounds are greater today than during the peak period of World War 11. Theoretically, several processes will increase the future supply of liquid and gaseous fuels, but the practical application of these methods depends on two controlling factors: 1—that enough carbon-bearing material be available to warrant industrial exploitation, and 2—that the given process be physically, chemically, and economically applicable. With regard to the first factor, the Western Hemisphere is fortunately well supplied, see Table I. Table I not only shows the extensive occurrence of many fuel sources, but also the wide discrepancy among quantities of fuel substances in various materials. Great quantities of potentially available fuel have not been recovered. Many coal seams and other hydrocarbon-bearing deposits are so situated that production is uneconomical by present-day mining or oilwell pumping methods, and even in many exploitable deposits where the most efficient production methods are used a large proportion of material is never recovered. Industrial production of hydrocarbon liquids and gases must be better balanced, and processes must be developed which will convert coals, oil sand, and oil shales into fabricated commodities. Among the techniques employed by scientists and engineers to secure fuel masses not economically available at present are several processes of underground gasification. These are all concerned with converting raw materials into gaseous products by thermal treatment of source substances while they are still in the ground and therefore require that a subsurface heat (fire) zone be induced to affect pervasively the buried raw fuel deposit. This is possible only where sufficiently high temperature exists and where the fuel material is satisfactorily opened or channelized. Also, to compete successfully with common mining or oilfield practices, a given method must be controllable and must not involve greater operational expense than regular methods. Preferably, it should be less expensive. Since the close of World War 11, England has continued field studies of underground gasification methods. Experiments have also been conducted in Sweden, Estonia, France, Belgium, Italy, and Germany. The U. S. Government has sponsored field studies at Gorgas, Ala., and because gasification in situ is vitally important to establishment of the nation's fuel resources, the mining engineering department at the Missouri School of Mines has been actively interested in developing the process of underground electrocarbonization. Research was conceived by Erich Sarapuu, and preliminary experiments in treating bituminous coal were begun at the school in September 1947. The work consisted of laboratory experimentation until June 1948, at which time, through a cooperative agreement with the Sinclair Coal Co. of Missouri, pilot plant studies were instituted at the Tiger mine of the Hume-Sinclair Coal Co. near Hume, Mo., to test the encouraging laboratory findings. This comprehensive field program was pursued continuously until the beginning of 1952."'" The broad term underground electrocarbonization describes a unique gasification process that releases by electricity the desired economic constituents from a raw carbon-bearing material, converting the substance of such deposits into deliverable form and greatly reducing the cost-of-mining factor. Underground electrocarbonization is not subject to the major limitations of many other underground gasifi-

Jan 1, 1955

By V. F. Parry

The major industrial uses of char or semicoke produced by low-temperature carbonization of coal4/ are as fuel for generating electric power, for blending with metallurgical coking coal as a substitute for low-volatile coals, and for gasification to produce synthesis gas for chemical purposes. This paper discusses various methods for carbonizing coal, emphasizing processing of the noncoking coals to produce char for thermal power-plant fuel and for blending with metallurgical coking coals and tar for chemical use. Low-temperature carbonization is a relatively simple processing step involving heating coal to about 900° to 1,200° F. to decompose the coal into 4 main products: (1) Semicoke or char; (2) complex, partly oxygenated liquid hydrocarbons or tar; (3) fixed gases; and (4) water. If the coal is heated rapidly under conditions that prevent decomposition of the primary products, 68 to 87 percent of the net heat in the raw coal is left in the char and 10 to 23 percent of the potential heat appears in the primary tar. Fixed gases represent approximately 7 percent of the heat and approximately 2 percent is lost or dissipated in the reactions of decomposition. Considering the different ranks of coal in the United States, the yield of tar will vary from about 9 to 44 gallons per ton, therefore it is important that each coal be carefully considered if it is to be carbonized for industrial use.

Jan 1, 1955

By R. W. Jenkins

RACKLESS equipment is being used by the T Coplay Cement Manufacturing Co. to develop a folded limestone deposit economically and safely with inexperienced men. Cost and quality of development stone results in a saving when it is used instead of outside sources of limestone. This favorable cost has been attained by following a methods development and training program described in this paper. It is expected that further savings in favor of the mined stone will be made when stopes now being prepared are in production. The present equipment will be used for loading from stope draw points. The limestone is blended with quarried cement rock to serve as raw material for Portland cement, first produced in this country by Coplay Cement Manufacturing Co. from a process developed by its president, David 0. Saylor, in 1873. For many years the cement rock alone or with a small amount of added limestone was used for raw materials. Later as specifications were raised and the richer portions of quarries were depleted, limestone from outside sources was required in increasing amounts. In 1925 D. J. Uhle, the general manager, began to prospect company property for limestone. It was known that the argillaceous cement rock was underlain by crystalline high calcium limestone. After an extensive drilling program it was found that there were large reserves of limestone underlying a recumbent syncline exposed in the south wall of the company's No. 3 cement rock quarry. A quarry was started on the vertical limb of this fold in 1927, and by 1930 the full requirement of the plant was met. After 1945 production diminished because of the occurrence of dolomite beds in the south wall. The quarry floor was then 135 ft below the cement rock level and 285 ft below surface. In 1949 Gerald I?.

Jan 1, 1955

By John L. Baum

Tests show that 50 times the earth's normal field can exist near stacked drill rods. Protection against the effect of these strong fields can be obtained by means of a removable sleeve of common iron pipe. GEOLOGISTS and engineers working around a diamond drill rig have often had the opportunity to observe the magnetic attraction of drill rods pulled out of the hole for core removal. This is sometimes an amusing pastime, but this same magnetic force inadvertently applied to a compass needle can cause it to lose its polarity entirely or to become reverse polarized when placed in the immediate vicinity of stacked rods. The first recommendation, of course, would be to avoid the rods, but in surveying an operating drillhole this is not always possible; therefore steps must be taken to neutralize the magnetic effect of the drill rods. Early in the summer of 1949 the geological department of the New Jersey Zinc Co. undertook at Franklin, N. J., an extensive program of surveying the course of deep diamond drillholes. The majority of the holes were surveyed while in the process of being drilled, and considerable difficulty was encountered. A pattern for the deviation of drillholes had been established for the district from surveys in holes from which the drill rig had been removed. It was observed that a certain operating deep drillhole being surveyed did not fit this pattern. In the resulting confusion of repeated surveys and cementing a plug for wedging operations, the lower 1800 ft of the hole was lost. Because of this incident studies were initiated to determine the source of the difficulties and to develop safeguards against recurrence of faulty surveys. Using a hand magnet easily shows that the polarity of a compass needle can be reversed if rotation of the needle is hindered through tilting of the compass or clamping of the mechanism. Such experiments are not recommended, especially in the case of the clockwork type of instruments. A magnetized hairspring in a clockwork instrument will cause the compass card to assume a periodic motion which will not stop until the clockwork runs down. Routine check readings of instrument north before and after each survey are not always sufficient to catch instances of reversed polarity of the compass needle. Although the majority of instances probably take place while the instrument is being removed from the hole with needle in fixed position, it is demonstrable that a 'luggish needle, weakened during a previous test, can be reverse polarized

Jan 1, 1955

By O. N. Carlson, H. A. Wilhelm, J. M. Dickinson

On the basis of microscopic studies, melting-point observations, and X-ray analyses, a phase diagram is proposed for the Cb-V system. A complete series of solid solutions is formed with a minimum in the solidus at 1810°C near 35 wt pct Cb. No compounds or intermediate phases were found in the system above 650°C. THERE is an ever increasing need for better structural metals and alloys for use in nuclear reactors. In addition to the normal properties of engineering structural materials, such as high temperature strength, resistance to corrosion, and fabric-abil~ty, the nuclear properties of the material must be considered. In a nuclear reactor it is important to conserve neutrons, so a material which removes these neutrons from the reaction excessively is considered to have unfavorable nuclear properties. In nuclear-reactor design the engineer must have nuclear as well as other data available on alloys in order to make a wise selection of materials. Due to the fact that many of the common structural materials have undesirable nuclear properties, it is vital that new alloys of metals having more favorable nuclear properties be investigated. Columbium and vanadium are both high melting metals, both exhibit resistance to chemical attack, and no great difficulty is encountered in fabricating them into desired shapes. With proper treatment both metals can be cold rolled extensively without failure. In addition they have desirable nuclear properties for certain types of reactors. Therefore, the alloys of columbium and vanadium should be of interest in the atomic energy program. Since an alloy-development program is enhanced by a knowledge of the phase equilibria of the components, this investigation was undertaken to establish the phase diagram for the Cb-V system. According to the Hume-Rothery rules for alloying,' the chemical similarity, crystal structure, and atomic-size factor are favorable for a complete series of solid solutions for this system. Both elements are in the same family of group V of the periodic table and thus are quite similar chemically. The crystal structures of columbium and vanadium are compatible for extensive solid solubility, since both have body-centered-cubic structures. The atomic diameters of columbium and vanadium are 2.85 and 2.62Å, respectively. This difference of slightly more than 8 pct is well within the 15 pct maximum difference allowed for extensive solid solubility. Experimental Procedures Source of Materials: Columbium powder and sheet trimmings were obtained from the Fansteel Metallurgical Corp. According to the manufacturer the metal contains less than 1 pct impurity. An analysis of the metal showed approximately 1800 ppm C in the powder while the sheet trimmings contained less than 500 ppm C. Spectrographic analysis showed minor amounts of Ca, Cr, Fe, Mn, Si, Ti, V, and Zr in both forms of the columbium. No commercial source of vanadium having the ductility and purity desired was available to the authors at the beginning of this investigation. As a result, all of the vanadium used in this study was prepared by the bomb reduction of vanadium pen-toxide with calcium employing the method reported by Long.' Yields of massive vanadium normally were about 80 pct. Chemical analysis of the vanadium prepared in this manner showed the presence of 200 to 500 ppm N and 800 to 1000 ppm C. Minor amounts of Ca, Fe, Mn, Si, Zr, Cr, and Cb were detected by spectrographic analysis. This vanadium metal was ductile and was cold rolled into 5 mil sheet. Annealing was not necessary during this rolling and the metal retained its cold-rolling characteristic after are-melting. Preparation of Alloys: The Cb-V alloys were prepared by melting pieces of vanadium sheet togethel-with columbium in the form of sheet or pellets of powder. The melting was carried out under argon in conventional arc-melting equipment employing a tungsten electrode and a water-cooled copper crucible. Each alloy was remelted three or four times, inverting the alloy after each melting in order to assure complete mixing. Alloys normally were obtained as round flat disks, weighed approximately 70 grams, and had roughly the shape of a disk 1 1/2 in. in diameter and 1/4 in. thick. Half of each alloy

Jan 1, 1955

By F. W. Wood

Ingots of zirconium and other highly reactive metals such as titanium and hafnium produced by the consumable-electrode arc-melting process require conditioning before forging or rolling. A nondestructive, arc-conditioning technique has been developed as a substitute for lathe conditioning. The technique employs en electric arc under an inert atmosphere to fuse surface metal on the sidewalls of ingots without seriously affecting the general quality of the ingots. Advantages of the practice are Increased yields, lower accumulation of inferior scrap, and actual cost savings. The work cited in this report was done on zirconium. INTRODUCTION The consumable-electrode arc-melting process as practiced at the Bureau of Mines3 4/ has proved to be a technique for producing sound, homogeneous, commercial-size ingots of zirconium and its alloys. Ingots produced by this process, however, have as a characteristic a layer of partly fused metal on the sidewall to a depth of 3/16 to 3/8 inch. To prevent extension of surface imperfections during forging, the ingots are ordinarily conditioned by removing defective material in a lathe. If the metal were less reactive or the requirements less severe, the machine chips could be recycled and remelted. However, zirconium, being so oxygen sensitive, is contaminated by merely being machined despite extensive cooling of the tool bit. Further, being contaminated with oxygen, the chip metal upon remelting is 10 to 25 Brinell-hardnese numbers harder than the ingot from which it was machined. For this reason, the ingot specifications for naval reactor use have prohibited remelting machine chips, except as required for starting pads. Again, were zirconium less costly or less in demand this ever-increasing pile could be burned to oxide and recycled through the Kroll sponge process. Economics do not favor this, as melting feed stock has an estimated increase In value over oxide of about $8.80 per pound of contained zirconium. The efforts of the Bureau to solve the problem have been directed along two lines: To reduce the contamination of the machine chips produced and to reduce the quantity of the chips produced. This paper presents the technique developed to reduce the quantity of chips produced by arc-conditioning the remelted ingot.

Jan 1, 1955

By H. H. Waller, D. L. Monroe, D. P. R. Smyth

ORIGINALLY slusher drift development was conventional, advancing the drift full 10x13-ft size at 6 ft per round. This proved dangerous and costly because the weak fractured rock of the orebody cannot be left raw for any length of time without its reacting with air to cause spalling. With this method of drift development ground conditions were worsened by many blasts and the length of time it was necessary for the drift to remain open. The problem was practically solved by driving 5x7-ft pilot drifts, then enlarging to 10x13-ft raw size by slashing with 50 to 60-ft diamond drill holes. Calculated explosive charges are used to avoid wall and roof shattering. Still more recently jacklegs and sectional steel replaced the diamond drills, resulting in the same work for less than half the cost, see Figs. 1 and 2. Crosscuts at right angles to the main haulage drifts, which cross through the center of the ore-body, divide the ore into 200-ft blocks. Eight slusher drifts on 50-ft centers draw from one block. The slusher floor is 11 ft above the rail, inclined at 1 pct grade, and the top is 18 ft below the undercut. The finished slusher drift is 102.5 or 120 ft long, depending on whether it is north or south of the crosscut. Drawpoints are staggered on each side at 17.5-ft centers. The completed drift is concreted and has an inclined ramp and a vent at the back. When crosscuts and gangways have been excavated and concreted and slusher stations installed, slusher drift development is carried out in 10 clear-cut stages: 1—Pilot drift is driven full length, 5x7 ft. 2—Station is excavated at half-way point in pilot drift. 3—Pilot drift is enlarged to slusher drift size by drilling. 4—Lower part of slash is blasted and mucked for floor concrete. 5—Floor is poured. 6—Inside upper half of slusher drift is blasted, mucked, and concreted. 7—Outside upper half of slusher drift is blasted, mucked, and concreted. 8—Drawpoints are excavated. 9—Drawpoints are concreted. 10—Concrete ramp and sheave protection recess are constructed at back end of slusher drift. A 5x7-ft pilot drift is driven full length on the center line of the proposed slusher drift using a jackleg and detachable bits on 7/8-in. hexagon steel. A 10-hp air-driven slusher and 36-in. scraper pull muck directly into cars in the main haulage crosscut, see Fig. 2. Operations are on a three-shift basis. One man mucks the previous round, drills, and blasts his round in a shift. Average advance is 5 ft per shift. When the pilot drift is completed, 7 ft of drift in the center section is enlarged to full raw slusher drift size of 10x13 ft, for the set-up to drill long slash holes. The pilot drift is slashed to full raw size from this station with jacklegs and tungsten carbide chisel bits on 7/8-in. sectional steel. Crews line the holes without trouble, working from controls established by the engineers, and the pattern of holes for the slash is arranged to give the least shattering effect

Jan 1, 1955