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Report of President The important fact of the year 1916 is that on July 25 contracts were executed by which the American Society of Civil Engineers because an additional Founder Society and arranged to make its permanent home in our building. The contracts provide for the construction of three additional stories to the building, the American Society to pay the cost tip to $250,000, this being considered equivalent to what each other Founder Society had contributed, the American Society to participate in the original Carnegie gift and have an equal share in the property with each other Founder society. The construction of the addition to the building is under way, $62,525.04 having been expended thereon in 1916. It has been found that, clue to war conditions and the advanced price of materials, the cost of the addition will be $300,000. The founder societies have been requested to assume the extra $50,000, one quarter each. This work is in charge of a Building Committee consisting of H. H. Barnes, Jr., Chairman, E. G. Spilsbury, Charles Warren Hunt and Charles F. Rand. At the request of the Founder Societies, important alterations were made in the lecture halls on the fifth floor of the building to make the same suitable for social functions of the societies. The cost, $6,642.08, was advanced by the United Engineering Society but is to be paid eventually by the Founders. The Library of the civil Engineers is being merged with the Library of the United Engineering. Society and the other founder societies. At the present time the membership of the four founder societies is 29,000 and of associate societies 23,000, so that a total of 52,000 engineers now have their headquarters in our building. The value of the real estate now owned by this Society is $1,647,171.16. This sum will be increased at the end of 1917 by the amount of the cost of the addition to the building.

Jan 3, 1917

By J. Downie, H. A. Kendall, G. H. F. Gardner

Some cases of the motion of two miscible fluids in uniform linear models are discussed. There is no bulk flow through the models, and the convection currents are caused solely by density gradients. Horizontal, vertical and inclined positions of the models are treated separately. Analytical formulas for the motion of a narrow transition zone between the fluids are obtained and compared with measurements, and the limitations are discussed. Observations of the ultimate stabilizing effect of diffusion on the motion me compared with the theory. It is shown that fluid motion in an inclined model with a constant rate of bulk flow can be directly deduced from the cases discussed, provided the viscosity ratio is small. INTRODUCTION In some reservoirs, convection currents caused by variations in the fluid densities may be significant. For practical purposes it may be desirable either to minimize the effect, particularly in miscible displacement, or to maximize the effect as in some schemes for the production of heavy crude oils by combustion. Experiments by Craig, et al,l have shown the effectiveness of gravity in reducing the efficiency of a miscible displacement when the mixed zone between the displacing and displaced fluids is narrow compared with the height of the system. The presence of a board transition zone from one fluid to the other has been shown by Perkins, et al,2 to nullify considerably the influence of a difference in density. Indeed, the theoretical discussion given by Perrine3 has shown that when the transition zone exceeds a certain critical length the efficiency approaches 100 per cent. Many additional results 495 have been published on miscible and immiscible displacements which show the importance of density difference and also mobility ratio. The aim of the present paper is to discuss some cases of gravity segregation of miscible fluids in linear models such as are often used in the laboratory to evaluate recovery processes, and to supply additional experimental evidence in support of the conclusions. An attempt is made to calculate simply the magnitude of the fluid motion caused by density pdients and mobility ratios, and to quantify the modifications in the fluid motion produced by the combined effect of molecular and convective dispersion. However, it should be pointed out that the uniform permeability of the models and the use of only a single pair of fluids (whereas several pairs of fluids may be employed in a recovery scheme) severely restricts direct conclusions about natural reservoirs. For conciseness, the motion generated by density gradients is emphasized for the case of no bulk flow through the models. This is not a drastic limitation because, as shown in Appendix A, a fluid motion with a constant bulk flow through any linear model is equivalent to a fluid motion with no bulk flow in the same model inclined at a different angle to the horizontal, provided only that the density is a linear function of the viscosity of the fluid mixture. HORIZONTAL MODEL In this case the model is assumed to have a rectangular cross-section with a horizontal breadth b, vertical height and a great length L. Throughout.the paper it will be assumed that x-axis points along the length, the y-axis along the width and the z-axis along the height. Thus, in the present case the x, y plane is horizontal. Initially, the transition zone between the fluids is assumed to be thin and parallel to the y, z plane. An example of this situation in a practical problem arises in a scheme for using air for cushion gas in the storage of fuel gas in aquifers. 6 With the passage of time the denser fluid sinks and spreads along the bottom of the model while the lighter fluid rises. The interface extends from the top plane of the model to the bottom. Experiments have shown that for practical purposes this interface may be treated as a plane surface at all times. The surface is significantly curved, however, for viscosity ratios greater than about 10. Other complications arise with large viscosity ratios because of the nonlinear dependence of the viscosity on the composition of the mixture, so that the simple results given here are restricted to fairly low viscosity ratios.

Jan 1, 1941

By S. F. Shaw

The Sierra Mojada mining district is situated in western Coahuila, about 8 km. east of the Chihuahua state line, and about one-half way from the northern to the southern extremity of the state. It is at the terminus of the Mexican Northern R. R., 125 km. northeast of Escal6n, Chihuahua; also about, 240 km. due west of Monclova, Coahuila. The altitude is approximately 1500 m. above sea level. The Sierra Mojada valley is about 10 km. long, from east to west, with a width of 1 to 2 km. It is approached from the east through the Puerto del Oro, where the Mexican Northern R. R. enters the valley; the slope from west to east is gentle. The Sierra Mojada, of which about 300 m. is an almost vertical wall, bounds the valley on the south side; on the north side, the slope to the crest of Sierra Planchada is more gradual, rising only 200 or 300 m. above the valley. Water is relatively scarce in the district, and there are no running streams. The rainy season, usually beginning at the middle of June, lasts through July, and occasionally into August or September. In some years the rains fail entirely. The slope of the ground is such that nearly all the water immediately runs off. The arroyos along the flanks of the mountains are not very deep, but in heavy rains, sometimes approaching cloudbursts, a tremendous volume of water rushes down them, carrying large boulders. A certain amount of seepage collects in the agglomerate, lying below the conglomerate, and is retained in the local fractures. Some shallow wells have been sunk in this agglomerate, and one well over 150 m. deep was sunk on the Volcan ranch, supplying about 76 1. (20 gal.) per min. A considerable supply of water is collected in large earth reservoirs in and below the town of Esmeralda. In ordinary years this supply will last until January or February, after which the Volcan well has to be depended on until the next rainy season. Vegetation is scarce, aside from scanty scrub oak and other brush in the mountains, and a few mezquite, locust, pepper, and cottonwood trees

Jan 1, 1923

By S. F. Shaw

The Sierra Mojada mining district is situated in western Coahuila, about 8 km. east of the Chihuahua state line, and about one-half way from the northern to the southern extremity of the state. It is at the terminus of the Mexican Northern R. R., 125 km. northeast of Escal6n, Chihuahua; also about, 240 km. due west of Monclova, Coahuila. The altitude is approximately 1500 m. above sea level. The Sierra Mojada valley is about 10 km. long, from east to west, with a width of 1 to 2 km. It is approached from the east through the Puerto del Oro, where the Mexican Northern R. R. enters the valley; the slope from west to east is gentle. The Sierra Mojada, of which about 300 m. is an almost vertical wall, bounds the valley on the south side; on the north side, the slope to the crest of Sierra Planchada is more gradual, rising only 200 or 300 m. above the valley. Water is relatively scarce in the district, and there are no running streams. The rainy season, usually beginning at the middle of June, lasts through July, and occasionally into August or September. In some years the rains fail entirely. The slope of the ground is such that nearly all the water immediately runs off. The arroyos along the flanks of the mountains are not very deep, but in heavy rains, sometimes approaching cloudbursts, a tremendous volume of water rushes down them, carrying large boulders. A certain amount of seepage collects in the agglomerate, lying below the conglomerate, and is retained in the local fractures. Some shallow wells have been sunk in this agglomerate, and one well over 150 m. deep was sunk on the Volcan ranch, supplying about 76 1. (20 gal.) per min. A considerable supply of water is collected in large earth reservoirs in and below the town of Esmeralda. In ordinary years this supply will last until January or February, after which the Volcan well has to be depended on until the next rainy season. Vegetation is scarce, aside from scanty scrub oak and other brush in the mountains, and a few mezquite, locust, pepper, and cottonwood trees

Jan 1, 1923

The following is an incomplete list of members and guests who called at Institute headquarters during the period Feb. 10, 1919, to Mar. 10, 1919. R. H. Allport, Cleveland, Ohio. R. R. Landon, Philippine Islands. I. H. Bergstein, New York. Alfred 0. Law, Cambridge, Mass. Capt. R. S. Burdette, San Antonio, Tex. James M. Little, Towanda, Pa. William R. Chedsey, State College, Pa. F. E. Marcy, Salt Lake City, Utah. L. R. Clapp, Philadelphia, Pa. H. F. Nash, Bartlesville, Okla. J. Murray Clark. L. S. Panyity, Columbus, Ohio. E. H. Claussen, San Francisco, Calif. W. C. Ralston, Reno, Nev. George E. Collins, -Denver, Colo. T. A. Rickard, San Francisco, Calif. George L. Collord, Sharpsville, Pa.. W. A. Rigby, Mt, Vernon, Iowa. William M. Corse, Mansfield, Ohio. H. H. Rowatt, Ottawa. Frank R. Crampton, Los Angeles, Calif. P. H. Royster, Pittsburgh, Pa. Capt. E. 0. Daue. Alan B. Sanger, Dorchester, Mass. A. E. Eddy, Butte, Mont. J. S. Shaw, Pittsburgh, Pa. N. H. Emmons, 2d, Lynchburg, Va. S. B. Shutts, Lynchburg, Va. E. W. Engelmann, Hayden, Ariz. Henry B. Smith, New Britain, Conn. F. A. Fahrenwald, Cleveland, Ohio. W. C. Smith, Roselle, N. J. 0. S. Finnie, Ottawa. Harry M. Spoor. John T. Fuller, Honesdale, Pa. D. M. Stackhouse, Johnstown, Pa. Major C. C. Griggs, Eureka, Utah. F. D. Stay,. Cleveland, Ohio. H. F. E. Gamm, Rutherford, N. J. D. B. Sterrett, St. Cyr, Que. A. W. Hackwood, Butte, Mont. R. H. Thayer, Yonkers, N. Y. C. L. Horne, Roanoke, Va. Kirby Thomas,. New York. B. B. Hood, Chrome, N. J. Henry Traphagen, Toledo, Ohio. R. E. Hore, Toronto Ont. H. D. Trounce, San Diego, Calif. Ross B. Hoffmann, Oakland, Calif. Henry J. Volkin, Ossining, N. Y. G. A. Joslin, Salt Lake City, Utah.. A. T. Ward, Bellefonte, Pa. John Johnston, Washington, D. C. Major L. W. Wickes. Paul S. King, Wilmington, Del. Horace V. Winchell, Minneapolis, Minn. E. William Kohl, Jr., Philadelphia, Pa. Fred L. Wolf, Mansfield, Ohio. W. H. Landers, San Francisco; Calif. J. Carson Adkerson, of the Stockwood Realty Corpn., Inc., of Woodstock, Va., writes that they are discontinuing their operations in Virginia on account of the conditions in the manganese market following the end of the war.

Jan 4, 1919

By Leonard Larson

BEFORE direct or wet smelting of copper concentrate was adopted at the McGill smelter, in November 1932, actual furnace smelting tests had indicated the possibility of smelting between 400 and 500 dry tons of wet concentrate per furnace day. Under calcine smelting practice, solid charge up to 1165 tons per furnace day had been smelted. It was realized that in order to smelt wet tonnage in equivalent amount per furnace day, a higher combustion rate and higher temperatures would be neces-sary within the furnace. The reverberatory furnace considered in the following history of wet-smelting de-velopments is a unit 127 ft. long and 28 ft. 8 in. wide, inside dimensions (Fig. r). This furnace was in operation through all the years from 1934 to date. During 1933, the first complete year of direct or wet smelting, the plant was on a curtailed copper-production basis and the available concentrate tonnage to be smelted was limited. The average coal-firing rate in the reverberatory for that year was 124 tons per furnace day. This rate was lower than the average for years immediately preced-ing under calcine smelting practice. With this coal-firing rate, an average of 391 tons of new metal-bearing material (N.M.B.M.) was smelted per furnace day. The following year, 1934, the coal-firing rate was in-creased to 139 tons and the smelting rate increased to 448 tons N.M.B.M. per fur-nace day. In succeeding years production required that either more tonnage be smelted in the one furnace or an additional furnace be placed in operation. With this situation confronting the smelting depart-ment, it was determined to continue the program of increasing the coal-firing rate and also to intensify the temperature in the smelting zone of the furnace. In order to accomplish this, various improvements were made to aid combustion within the furnace. The burners were enlarged, in-creased primary air was supplied to the burners, and changes were made in the furnace to enlarge and streamline the out-let, thus increasing the furnace draft.

Jan 1, 1940

By John E. Willson, Ben L. Seegmiller

A portable electronic instrument was designed and constructed to detect unknown stress magnitudes in rocks. The principle used to detect stress is based on the propagation velocity method. This method allows the stress in a rock subjected to a load of unknown magnitude to be determined, provided that the wave velocity-stress relationship for similar rock is known. This is accomplished by measuring the travel time of longitudinal mechanical waves passing between two points a set distance apart in a rock. The velocity of the wave is calculated and the stress determined from the wave velocity-stress relationship. If the sending and receiving transducer spacing is constant, a time vs. stress relationship rather than the velocity-stress relationship may be used. The method is nondestructive and tests can be made without drilling or otherwise disturbing the rocks. The first studies undertaken in the United States to determine stress in rocks using propagation velocity techniques were reported by Obert1,2 in 1939 and 1940. The Soviet Union first reported using propagation methods to study rock pressures in 1951.8 Success of the method led to the development of a pulse-type ultrasonic seismoscope4 in 1953. Using this instrument, Ivanov and Betaneli5,6 in 1963 succeeded in devising and testing under field conditions a method of investigating coal pillar stresses. In 1967 Osipova7 reported results of similar studies in the Nakhichevan salt mines. In France, Tincelin8 has used the propagation velocity method to study the stress distribution in iron ore pillars. Uhlmann9 has investigated stresses in salt and potash pillars in Germany, using velocity techniques. Determination of uniaxial stress by propagation velocity methods is limited to rocks which have readily detectable wave velocity-stress variations. The present stress detection instrument is restricted in application to rocks which have a velocity change under stress of at least 300 fps. Examples of rocks which meet this requirement are sandstones, coal, and possibly some limestones. Testing of this instrument was limited to a laboratory study and the results may or may not be indicative of what would be found in a field test. A program of field experiments to study the feasibility of using this instrument to determine mine pillar and tunnel stresses is in progress. Instrument Design The instrument has two main components: A probe and a control-display unit. The probe is a hand-held device to which two identical rodlike transducers are rigidly mounted. Coaxial cable connects the probe to the control-display unit which is mounted in an enclosed carrying case measuring 7 % x 9 x 13 in. The instrument is designed so it may be carried and operated by one man. Weight of the probe and enclosed control-display unit is approximately 20.1b. The probe consists of two transducers identical in construction. One transducer is used to convert electronic pulses to mechanical waves and transmit these waves into a rock. The other detects the transmitted waves in the rock and converts them back into electronic signals. The basic element in both transducers is a piezoelectric crystal. The crystal is a disk made of lead titanate zirconate and has a natural resonant frequency of 400 kHz ±1% in the thickness mode. A schematic of one of the two identical transducing elements is shown in [Fig. 1.] A spacing of 6 in. between sending and receiving transducers has been found to be most satisfactory. The control-display unit consists of a pulse generator and a 1-in. oscilloscope. Various electronic devices are used for the power supply, amplification, and calibration. The amplitude of the square wave from the pulse generator can be continuously varied between 0 and +20 vdc. Pulse width can be set at 1, 2, 3, or 4 µsec. The repetition rate of the pulse can have the following values: 50, 75, 100, 150, 200, 250, 300, 400, 500, 750, and 1000 Hz. The oscilloscope delay system allows the travel time of the longitudinal wave passing through a rock to be measured to an accuracy ±0.1 µsec in the range from 0 to 995 µsec. Fig. 2 is a schematic of the various electronic sections in the control-display unit. Laboratory Testing The first step in using the instrument is to develop a velocity-stress or a time-stress curve for the par-

Jan 1, 1972

By W. C. McClain

In the design of underground excavations, rock mechanics considerations are nearly always based on an elastic behavior of rock. Most rocks do exhibit a certain amount of elasticity, and the application of principles and concepts of elasticity to rock mechanics problems has been largely responsible for progress in the field over the past 10 or 15 years. No doubt the theory of elasticity will remain the most useful analytical tool available for both the design of underground openings and the interpretation of phenomenological data.

Jan 1, 1967

By O. Hall, V. P. Row

Diamond drilling has become a very important factor in mining. It is the most satisfactory method of obtaining proof of the existence of an orebody and of determining the character and extent of the body in measures where a core can be made. Large investments are based on the information obtained. The method of development, the mine and reduction plants, and subsidiary essentials established are often without other information. The Mond Nickel Co. has drilled about 50,000 ft. (15,240 m.) per yr. for sevéral years and has given considerable study to the question of improving the methods and of overcoming the limitations of this means of securing information. One of the main limitations of the diamond drill has been its inability to drill a straight hole. To depths of 1000 ft. (304 m.) the deflection is often not large but all holes tend to curve and take somewhat erratic courses. A hole running at a small angle to a fissure plane tends to run along the fissure. If the hole is at a large angle, the tendency is to penetrate the fissure plane at right angles. The face of the bit having one side in solid formation beyond the fissure, the other in the selvage of the fissure, tends to turn, the rods curving to the extent of their clearance in the hole. Usually rock measures have somewhat regular fissuring or strata and holes take somewhat similar courses, usually at right angles to the fissuring or strata. Excessive pressure or short or worn core barrels increase curvature. Slow speeds, correct feeds, sharp bits, and core barrels the full size of the bit, tend to decrease curvature but do not prevent it Curvature once started increases rapidly. The disappointments from crooked holes have been numerous. Discoveries have later been found to be on the properties of others. One Canadian company had a hole, supposedly discovering ore 300 ft. (91 m.) from old workings, unexpectedly turn in 8 ft. (2.4 m.) of wooden core from a timber in the old workings. In many cases the information secured by a curved hole is satisfactory, if it can be accurately surveyed. The inclination of the hole at any point

Jan 1, 1920

By W. W. Stephen, J. P. Walsted, W. J. Kroll

FOR the production of carbides, various furnace types are available, especially those using arc, resistance, and high-frequency heating. Selection of a specific means of heating depends primarily on the material to be treated and the physical properties of the carbide produced. In the present case, zirconium carbide had to be prepared on an industrial scale as a raw material for the production of anhydrous zirconium chloride. Considering that a rather expensive pure oxide was to be used, the arc-furnace treatment recommended for zircon sands in a previous publication' was ruled out because of the considerable volatilization and dust losses caused by the blast of the arc. For this reason, either high-frequency or resistance heating seemed to offer more promise. Since there was not enough capacity of the former available, resistance heating was chosen. It was first thought that the Acheson silicon carbide furnace would be suitable for the present purpose, but the voltage in such a furnace, in which the current passes through the batch, varies from 220 to 75 v from the start to the end of a run. This variation is so great that a special tap transformer would have been required. Trouble was also expected by local melting of the carbide. Pure zirconium carbide melts at about 3527°C, but the melting point is brought down to 2427°C, according to Agte,2 when an excess of 6 pct C is present in the carbide. This we found confirmed by experiments in a high-frequency furnace. Excess carbon is needed in the batch to obtain a complete reduction. Fusion of the charge would cause great difficulties in an Acheson-type furnace because of the good electrical conductivity of the carbide as compared with that of the loose batch. Also, fused carbide is much more difficult to chlorinate than the spongy product that can be made in the radiation furnace described below. It was apparent that, to obtain a good-qual-ity zirconium carbide, the heat input would have to be well-controlled. The hairpin-resistor principle seemed to offer possibilities in this regard, and a furnace of this type was therefore developed. The advantages of the hairpin-resistor radiation principle have been discussed in previous publications, and a split-tube graphite-resistor furnace," now increasingly used in various laboratories, as well as a centrifugal quartz melting furnace4 of this type, has demonstrated the usefulness of this heating method. The hairpin-heater element has the following definite advantages over a straight resistor of the type used, for instance, by Georges:5 Its resistance is four times greater; it can expand freely; it is sturdier because of the larger diameter, and it has a larger radiation surface; there are no hot contacts that might wear out or overheat; only one clamp is used which permits assembling all electrical leads at one side of the furnace, making the other sides easily accessible to the operator. The shorter element and its larger diameter permit greater concentration of heat. The furnace developed is shown in fig. 1. The box (I), made of 2 1/2-in. graphite plates, has inside dimensions of 23x17x16 in. It contains the briquet-ted batch (2). The box is embedded in lampblack (3) up to the cover plate. The cover plate contains an opening for the gas escape (4) and for the observation hole (5), which permits measuring the temperature with an optical pyrometer. The cover plate is embedded in charcoal (6). The lampblack is contained in the insulating brick lining (7), held in the 1/4-in. sheet steel box (8). The graphite box is set on two rows of triangular graphite bars (9). The hairpin-heater element (10), the dimensions of which are given below the main drawing, extends horizontally in the graphite chamber and radiates freely on the batch. A graphite tube (11) keeps the lampblack from falling into the slot. The split electrode, which in reality is turned 90" against the drawing, is so arranged that the slot is vertical. The water-cooled packing gland (12) is insulated by an airgap from the heater element. A thin pipe (13)

Jan 1, 1951

By Philip J. Shenon

IN 1947 the mining industry strove desperately to regain operating normalcy. During the first part of the year the industry in this country was plagued with labor shortages, strikes, and portal-to-portal lawsuits. Then followed the uncertainty of Government policy with respect to subsidies and tariffs; and the long and bitter controversy as to whether we are a nation of mineral "haves" or "have-nots." As the year progressed the labor supply impoved somewhat and the Taft-Hartley law brought partial relief to strike-troubled industry. The uncertainty with respect to tariffs was partly answered by President Truman on April 29 when he announced the suspension of the 4¢ import duty on copper until March 31, 1949, and later, near the close of the year, by the announcement of the tariff agreement signed by the United Nations at Geneva. By this agreement tariff rates are to be adjusted downward, or "bound" at their present levels, for three years, on commodities amounting to some 60 per cent of United States trade. Effective Jan. 1, 1948 the rate of duty will be substantially reduced on ores, concentrates, or metals of zinc, aluminum, copper, manganese, tungsten, iron, steel, barytes, nickel, cadmium, and antimony.

Jan 1, 1948

By William J. Robinson

What is the cumulative rate of recovery of copper from a sulfide leach dump? The technical answers to this frequently asked question may vary from "I don't know" to "quite good" from people of the same company. Management also poses this question because it plays a vital role in planning future production. The answer is difficult because of the many variables between different operations such as physical construction of the dump, topography of the original site, characteristics of dump material, mineralization, and water application. Each of these factors and others have a bearing on the rate of recovery. Anaconda's method of water distribution at Butte is unique in that each dump is drilled to allow physical injection of water into the dump. The injection holes are drilled vertically on 25-ft centers from the surface of the dump [(Figs. 1 and 2)]. Normally, the holes are drilled 40 ft deep and cased with 4-in. PVC perforated pipe. The pipe is perforated at 1-ft intervals with four 1 -in.-diam holes at 90° spacing on the pipe circumference. Formerly, shallow holes (20-40 ft) were alternated with deep holes (80-180 ft) in an effort to cover the entire dump and hasten oxidation. The shallow hole depth varied according to the compaction thickness. The deep hole depth varied according to the height of the dump and the original ground profile. It is preferable to have the deep holes end 20-30 ft above the original ground level. This generally allows penetration of the coarse material at the bottom of the dump. Deep holes had to be discontinued in order to cover more

Jan 1, 1974

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Jan 1, 1938

For the convenience of members who may be traveling from place to place, the Secretary has compiled a list of libraries or other places in the United States and foreign countries where MINING AND METALLURGY or the Transactions may be consulted. This list will be sent to members if they so request.

Jan 1, 1925

At its meeting on February 21, 1933, the Board of Directors passed the following resolution defining and expressing the policy of the Institute with respect to official participation or action in controversial matters: WHEREAS, the Board of Directors at its last meeting passed the following resolution: "The Board of Directors, having been appealed to by several Local Sections for advice concerning the attitude and policy of the Institute relating to the endorsement of candidates for public office, the following statement of

Jan 1, 1938

Jan 1, 1911

By E. S. Wheeler

THE conversion plant of the Climax 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, Colorado.1 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. o.d. 8-hearth Nichols-Herreshoff); also the first experimental ferromolybdenum was produced at Langeloth. At this time calcium 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. o.d. roaster was erected at this time (1929), plans were prepared for a duplicate unit, which was completed in 1935, 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. o.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. o.d. roasters. These two new units, along with a change in roasting practice, brought the potential roasting capacity to 150,000 to 190,000 lb. 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

In the examination of an undeveloped prospect a decision must be arrived at from an inspection of the outcrops and the exposures in a few shallow pits. Prospects that are offered for sale rarely expose any important quantity of good ore, work having usually been stopped when the immediate exploration no longer yielded favorable results. It should be borne in mind, furthermore, that a great majority of prospects have been examined many times, and if of conspicuous promise, would have been acquired-for development. In most cases the problem for the engineer, therefore, is to discover the traces of a valuable mineralization that has disappeared through solution, or the conditions that indicate the possibility, or probability, of underlying secondary enrichments. The available data being meager, every feature should be the subject of careful study-the condition of the outcrop, the structural relationships, the associations of minerals, the alterations of wall rocks, the chances for underlying enrichments, and so forth, as well as the actual assay value of the material exposed. It is frequently advisable to spend the time and money necessary to have trenches dug at various significant points along a promising outcrop; such exposures, even if the depth attained is slight, often disclose conditions that are not apparent at the actual surface, as for example, the existence of soft minerals and the distribution of the various minerals through the mass of the deposit. Trenches also permit samples to be taken from points that were not accessible during former examinations. Strong, persistent outcrops of uniform width may be taken to indicate the probable size and character of the underlying vein,

Jan 1, 1932

Truman H. Aldrich W. Lawrence Austin W. S. Ayres John Henry Banks Wrn de L Benedict Andrew A Blair Walter M. Brodie H. H. Burden Charles Butters

Jan 1, 1932