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By R. S. Handy

Does hand-sorting of mill feed pay? The fact that the practice is so general would seem to indicate that there must be good reasons for following it; yet, to my mind, the advantage in many cases is doubtful enough to invite a thorough discussion of the subject, in an effort to determine under what conditions hand-sorting does pay. A typical sorting plant in the Cœur d'Alene district handles about 800 tons of mine-run ore per day. The material is run over a grizzly with about 4-in. opening; the oversize is crushed to about 4-in. size and joins the undersize, and the whole product is washed and screened on trommels with about 11/2-in. openings. The oversize from these trommels is spread on a belt and the waste and first-class ore are sorted out by hand, the residue being further crushed, and passes, together with the trommel undersize, into the mill feed bins. About 50 tons of shipping ore and 150 tons of waste are sorted out per day, leaving about 600 tons of mill feed. To sort out this material requires 20 ore sorters and five bosses and repairmen. The normal cost of hand-sorting is about 16 c. per ton of run-of-mine ore, or 65 c. per ton of material sorted out. I believe that straight crushing and milling of the run-of-mine ore, in this case, is not only cheaper, but is metallurgi-cally more economical and efficient than hand-sorting followed by milling of the residue. At one of the Bunker Hill rock-houses, 1250 tons of run-of-mine ore are reduced to pass 30-mm., or about 11/4-in., opening in six hours, requiring five men. The equipment and the power consumed are not greatly different from that required in a sorting plant of the same capacity, and the normal cost of thus preparing the feed for milling is under 3 c. per ton of mine product. The relative cheapness of this method as against hand-sorting is thus clearly established, the saving being about 13 c. per ton. If there is any virtue in, hand-sorting, then, it must lie in the metallurgical results achieved. Possibly the favor that hand-sorting has found is due to a misconception as to what happens to the clean shipping ore if it is sent to the mill. If a mill makes, say, an 80 per cent. recovery on its feed, the hand-sorting advocate may assume that if the hand-sorted product be sent to the mill, 20 per cent. of the metals contained in it will be lost. This, I believe,

Jan 1, 1920

By John R. Green

ROTARY kilns, varying in construction, are used in a variety of processes. Products differ and operating conditions vary according to economic requirements. All of these variables influence the degree of instrumentation and type of control necessary or profitable to install. Products fired in rotary kilns are the results of combined temperature and time treatment and the interrelation of these two factors should be kept constantly in mind in the perusal of this paper. MEASUREMENTS OF KILN TEMPERATURE The securing of proper temperatures within the rotary kiln, either of the charge or of the kiln lining in the high fire zone, are basic objectives of kiln operation; therefore the measurement of these high-temperature conditions merits first consideration.. For these measurements, radiation pyrometers, properly designed and protected, have proved adequate, particularly where records or automatic control are desired, while optical pyrometers may serve as portable indicators. Light-sensitive devices or photoelectric cells in this service are capable of producing records as well as certain forms of automatic control. High-temperature measuring devices for rotary-kiln work should have the measuring unit protected from dust, as a continued accumulation of deposit would lower the readings. Fig. I shows a typical application of radiation equipment to a rotary kiln, mounted far enough away from the sighting hole through the kiln hood to be protected from puffs or stingouts and with a sighting tube having a series of settling chambers, to prevent dust from reaching the lens. Although dust conditions inside the kiln itself usually fluctuate considerably and have their effect upon the readings at that moment, such effects are not cumulative and show as a continuous fluctuation in the temperature indicated, provided the radiation unit and recorder are responsive to such slight and rapid temperature fluctuations. The ambient temperature to which a measuring unit is subjected fluctuates from possibly o° to over 200°F., therefore these changes should have no effect upon the final reading or upon the life of the measuring unit. A very narrow sighting angle is desirable, so that measurement of temperature may be made as far as possible into the high fire zone without. effect of the flame temperature. There should be some sighting means on the measuring unit, so that the proper field may be viewed. Measurement of wall temperature taken at a point where the charge has just left the wall is less affected by dusting conditions than when taken elsewhere. Factors Influencing Temperature Table I gives an outline of the factors that produce kiln temperatures and com¬ments on the possibility of measurement and control. Of greater number and of varied influ¬ence are the factors that control when, where and how the heat is delivered to the kiln, the type of flame and the rate of

Jan 1, 1941

By Blake M. Loring, Robert A. Colton

MELTING procedures for most metals and alloys usually include some provision for the control of oxygen, since this element frequently has some undesirable effect on the properties of the metal or alloy concerned. In the melting of the various bronzes, this is especially true, since oxygen can be introduced into the melt in a number of ways, most of which are difficult to control. In making bronzes from virgin materials, oxygen is usually present in the commercial grades of copper used; electrolytic he refined copper ingots usually contain from 0.02 to 0.05 per cent oxygen. When ingot metal or scrap bronze is used in the charge, oxides are almost always present to some degree. Under certain melting conditions and in certain types of melting equipment the atmosphere over the metal may be highly oxidizing. This, of course, may be desirable in the melting practice used, but is certain to introduce appreciable amounts of oxygen into the melt, and that may not be entirely desirable. From the foundry viewpoint oxygen may have several harmful effects on bronzes. The formation of excessive amounts of oxides may produce drossy metal, which, in turn, means dirty castings and a potentially high rejection rate. It is claimed, too, that oxides in the metal may combine with hydrogen to form steam and thereby produce porosity in certain of the bronzes. The presence of certain refractory oxides, such as SnO2, may be undesirable from the viewpoint of making machining difficult, or as possible stress-raisers in the metal matrix. It is often said that excessive oxide present in the liquid metal makes the metal run sluggishly. Since it appears impractical to try to control the amount of oxygen introduced into the metal beyond taking the economical limit of precaution about charging dirty scrap, it would seem that control of the oxygen content of the liquid metal offers the most promising approach to this problem. This might be done by using slags, or covers, or by the use of some element or compound added to the melt to remove oxygen preferentially. Most aspects of this problem have been considered and considerable work has been done along these lines. A reconsideration of the problem is indicated, however, by the controversial nature of the evidence at hand, especially on the possibility of using a chemical deoxidizer to control the oxygen content of bronzes. CHOICE OF DEOXIDIZER The choice of the chemical agent selected as a deoxidizer would naturally depend first on the alloy in which it is to be used; in this particular case the alloy is Navy gun metal. In Table I are listed a group of elements that might be considered as

Jan 1, 1946

By C. V. Millikan, Herbert F. Beardmore

Declining Productivity Index, as considered herein, is a productivity index that has a substantially and consistently decreasing value when measured over a period of a few hours. If not recognized, it may be interpreted as representing an unsettled or unstable condition in a producing well. It forms the basis of a number of predictions as to future performance compared with the performance of wells in which the productivity index does not decline. The nost important of these are that: (I) the rate of oil production will decline more rapidly, (2) the gas-oil ratio will usually increase abnormally, (3) the amount of water ultimately produced will be negligible, if any, (4) stop-cocking will increase current production and decrease gas-oil ratio. When such relative well performance is indicated, it follows that certain reservoir performance and operating practices may be anticipated: (I) disappointing recovery, (2) longer natural flow, (3) no active water drive, (4) better evaluation of acidizing, shooting and reworking, (5) intermittent operation of wells, (6) smaller pumping equipment required. It is significant that the basis of these important predictions is bottom-hole pressure tests that can be made immediately after completion of the wells. Introduction "'Productivity index" is the number of barrels of oil produced per day per pound per square inch of differential pressure between the reservoir and the bottom of the well. "Declining Productivity Index," as considered herein, is a productivity index that has a substantially and consistently decreasing value when measured over a period of a few hours. Productivity index can be determined only when the rate of production can be stabilized. In most wells the rate of production and bottom-hole pressure are easily stabilized to give a constant productivity index for a period of many hours. A few produce so erratically as to preclude probability of an acceptable test in the usual sense. A satisfactory productivity index can be determined in some wells only by gauging the production accurately by hours, measuring the bottom-hole pressure continuously, and from these data calculating the productivity index for successive hours. Such figures frequently will have a decreasing value—that is, a declining productivity index—and when plotted on double. logarithmic cross-section paper with productivity index as ordinate and time as abscissa will be a straight-line curve. The angle between the curve and the abscissa is referred to as the rate of decline. Declining productivity index may appear to be the result of change in permeability to oil due to decreasing saturation of oil and increasing saturation of gas in the reservoir. It is possible that such is in fact the cause, but if so, more is involved than merely relative saturation of oil and gas. This seems evident because the rate of decline in productivity index is substantially the same regardless of the pro-

Jan 1, 1945

By B. D. Cullity, S. S. Hsu

The stress distribution at the surface of a twisted cylinder is analyzed along the boundary of a slip plane of arbitrary orientation and this analysis is applied to the torsion of cylindrical crystals of magnesium. A criterion for slip is developed and the critical resolved shear stress of magnesium in torsion is found to be 0.038 kg per sq mm. The mechanism of torsional deformation in magnesium and aluminum is described, as well as the relaxation of a twisted crystal at elevated temperatures. THERE has been considerable fundamental study A of the mechanism of slip occurring during deformation in tension and compression but comparatively little attention has been paid to the mechanism of torsional deformation. This paper is concerned with the way in which a single crystal deforms in static torsion and with the phenomenon of untwisting which a twisted crystal undergoes during annealing. Particular attention is paid to the application of the critical resolved shear stress law to torsional deformation. Probably the most likely mode of torsional deformation that can be imagined consists of rotation of one part of the crystal relative to another on the active slip plane. Here the most interesting feature is the slip direction. If a cylindrical single crystal is twisted about its axis and if, to consider the simplest case, the active slip plane is normal to the specimen axis, then the direction of gross slip at any point must always be at right angles to the radius drawn to that point. It follows that the crystallographic direction of slip must be continually changing, in sharp contrast to slip in tension which occurs in a crystallographically constant direction. Extensive investigations of slip in torsion have been made by Gough and his associates.' They made torsional fatigue tests on metal single crystals belonging to a number of different crystal systems and observed, with a microscope, the appearance of slip lines on the surface and of strain markings on the transverse cross section of cylindrical specimens. They concluded that the critical resolved shear stress law was just as valid in torsion as in tension, in the sense that slip occurred on that plane and in that direction in which the resolved shear stress was a maximum. They also established that the operative slip planes and directions were the same as in tension. They did not, however, measure the critical resolved shear stress in torsion nor, in fact, did they establish that a critical value of the resolved shear stress was necessary for the initiation of slip. Their investigations were, moreover, restricted to torsional fatigue tests which necessarily involve small strain amplitudes. Wilman,' studying crystal growth and the abraded surfaces of crystals, observed crystal fragments

Jan 1, 1955

By A. H. Daane, B. J. Beaudry

The uranium-antimony system has been investigated by metal-lographic, therma1, and X-ray methods. There are four intermediate phases present: USbz and U,Sb, which undergo peritectic decomposition at 1355°and 1695OC, respectively, and USb and UPSbs which melt congruently at 1850"and 1800°C, respectively. There is a eutectic between U4Sbs and USb which melts at 1770°C. USb,, UsSb4, and USb are line compounds while U4Sbs has a solubility range of approximately 1.5 at. pct at 1770°C. Single crystal studies of the cmpo,und U.,SbS show it to have a hexagonal unit cell with a, = 9.268A and c, = 6.201b;. The solubility of uranium in liquid antimony increases from 0.1 at. pct at 650°C to 1.5 at. pct at 900°C. The solubility of antimony in a, P, and y uranium is approximately 0.02, 0.1, and 0.05 at. pct respectively. IN nuclear reactors, uranium has been used mainly in the form of solid-fuel elements of the pure-metal or uranium-rich alloys. However, in attempts to overcome some of the difficulties inherent in solid-fueled reactors, reactor concepts hive been proposed in which the fuel would be a liquid metal, such as a solution of uranium in another metal. The liquid-metal fueled reactor (LMFR) designed by the Brook-haven National Laboratory of the U.S. Atomic Energy Commission is an example. Fuel for this reactor is a solution of approximately 800 ppm enriched uranium in bismuth.' To design such liquid-fueled reactors, a background of information on the solubility of uranium in various metals is obviously necessary. The present study was undertaken to pro- vide information for such purposes, and in addition, to study the uranium-antimony system as a whole. A preliminary literature search showed that some work had been done previously on uranium-antimony alloys, but no attempts had been made to make a complete study of this system. The solubility of uranium in antimony was studied by Cammack and Bridger' up to 950°C and by Hayes and Gordon up to 900'-C. Their data are plotted in Fig. 1. Ferro, in an X-ray structure study of some uranium-antimony compounds, found USg to be tetragonal with a, = 4.2724 and c, = 8.471A, U3S4 to be b.c.c. with aO = 9.095A and USb to be f.c.c. with a,= 6.191A. Colani reported the compound UsSba; however, work by Ferro and by this Laboratory showed that this compound does not exist. foote' reported several intermediate phases in the uranium-antimony system, but did not identify them. Greninger' re-ported finding an intermediate phase in a 5 at. pct Sb

Jan 1, 1960

By Kenneth C. Vincent, A. M. Gaudin

In a study1 of the action of amyl xan-thate on chalcocite in water suspension, it was found that if the xanthate is added in relatively large quantity the xanthate group can be accounted for in three forms: (1) as xanthate ion remaining in water, (2) as benzene-soluble cuprous xanthate on the chalcocite, (3) as a benzene-insoluble coating on the chalcocite. Reduction in the amount of xanthate added leads to disappearance of form I then of form 2. With minute quantities of added xanthate, substantially all the reagent is of the "unleachable" kind. Yet this unleach-able xanthate can be dissolved by pyridine to form a soluble cuprous xanthate complex, and it is sufficient in effectiveness to make chalcocite flotable even if insufficient in quantity to form a complete mono-molecular film on the mineral. In that study it was also found that chalcocite abstracts cuprous xanthate from benzene so1ution, in what may (or may not) be a case of molecular adsorption. It seemed desirable, therefore, to study further the "unleachable" xanthate on the chalcocite, and particularly to ascertain whether it is formed as a result of the benzene leach or is actually formed during the conditioning of chalcocite in aqueous xanthate solution. A 'splendid tool to make such studies has recently been perfected; that is, the use of radioactive spies or tracer atoms. With this tool, it was thought possible to answer the problem stated above, but, unfortunately, this did not prove to be true. Answer to the flotation question was clouded by the existence of rapid and selective diffusion of copper atoms in the solid mineral—a phenomenon that was totally unexpected yet remains interesting and perhaps important. Method of Radioactive Spies The method of radioactive spies is discussed in many places. Two convenient references are the article by Evans2 and the book by pollard and Davidson.3 The method was first developed by van Hevesy and paneth in 1013, when only natural radioactive isotopes were available. These isotopes are few in number, and so the method remained limited in scope. Recently the synthesis of over 350 active isotopes of all the known stable chemical elements4 has opened widely expanded possibilities for the method, A fundamental tenet of the method is that a radioactive atom and a nonradio-active atom are exactly alike until disintegration of the radioactive atom. Then, by its sudden death, the spy reveals its location. The relative abundance of spies at various locations makes it possible to ascertain the relative abundance of the corresponding nonradioactive atoms at the same times and locations. A second fundamental tenet of the method of radioactive tracers is that the radioactivity lies in the nucleus of the atom: transfer of radioactivity from one

Jan 1, 1947

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the dose association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and' 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill 10,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. 1) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1017 and it was operated during parts of 1918 and 1910. In the postwar years, there was almost no demand for the product and production was at a standstill until 1024. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942

By Robert P. Burrows

Discussion of the paper of R. P. Burrows, presented `at the New York meeting, February, 1916, and printed in Bulletin No. 107, November, 1915, pp. 2237 to 2245. EDWIN M. CHANCE, Wilkes-Barre, Pa.-I have been very much interested in the comprehensive and able discussion of the important topic of mine lighting just given by Mr. Burrows, especially as I know by common report Mr. Burrows' ability and his close touch with the development of the electric mine lamp, and its use in mines in connection with a portable electric lamp. It is regrettable that Mr: Burrows was not able to acquaint himself more closely with the cost of mine illumination, the actual cost, as produced by a portable light as carried by miners, whether a flame safety lamp or an open light. Had he been able to acquaint himself more closely with these data, he perhaps would not have arrived at some of the conclusions which he has presented to us in his paper. For example, we find Mr. Burrows advises that the open torch is by far the cheapest source of illumination, as produced by a portable light. I am very sure that is not the fact. The cost of miners' oil varies from about 25.c. a gallon to 90b. a gallon. The various States have passed laws from time to time rendering it mandatory that the miner shall use, whether he is willing or not, a so-called better or higher grade of burning oil. These laws have entailed the use of an oil so expensive that the use of the oil torch has become practically impossible, and the acetylene or carbide miners' lamp has largely taken its place. I have evolved a few figures showing the relative efficiency and usefulness of the open oil torch and the acetylene miners' lamp. For example, the acetylene miners' lamp consumes or burns 4 liters of acetylene per hour, or 0.14 cu. ft.; the oxygen consumed is 10 liters, or 0.35 cu. ft. per hour; air consumed, 50 liters, or 1.76 cu. ft. per hour; black-damp produced, 40 liters, or 1.41 cu. ft. per hour; air rendered extinctive to the acetylene flame, 112 liters, or 3.95 cu. ft. per hour; carbide con- sumed, 13 g. per hour. The open oil torch, on the other hand, burns about 20 g. of oil per hour; the oxygen consumed is 48 liters, or 1.69 cu. ft. per. hour; air consumed, 240 liters, or 8.47 cu. ft. per hour; blackdamp produced, 191 liters, or 6.73 cu. ft. per hour; air rendered extinctive to oil flame is 1,413 liters, or 49.92 cu. ft. per hour.

Jan 5, 1916

By William M. Newton, John E. Sherborne

The relation between the properties of electric logs and the fluid used in drilling a well is briefly discussed in this paper, and the fact that the resistivity of the drilling mud and its filtrate can appreciably affect such logs is shown. Apparatus was designed and a procedure established to determine the effect of various factors on the resistivity of drilling mud. Five muds commonly employed in California were tested and the following observations were made: I. The effect of raising the temperature from 80°F. to 180°F. is to decrease the resis-tiviiy of the mud or filtrate approximately 50 per cent. 2. The resistivity of the mud in most cases closely approximates that of its filtrate. 3, The change in the resistivities of muds caused by the addition of chemical is not the same function of the amount of chemical added for each mud. 4. The effect produced by increased sodium chloride content is to reduce markedly the resistivity of the mud and its filtrate. 5. Weighting materials, such as Baroid and limestone, tend to increase the resistivity of drilling mud. 6, Cement and counteracting reagents reduce mud resistivity. Introduction During recent years, the practice of "electrically logging9' a geologic formation by means of a system of electrodes lowered in a well bore has provided data of inestimable value concerning the nature of the sediments penetrated. Thus, it is generally possible, by means of electric logs, to distinguish readily the differences between shales, sandstones, cherts, limestones, and other rock forms. The distinguishing characteristics recorded in this manner have made possible long-range and closely detailed mapping and visualization of deeply buried structures1-5 to a degree not to be obtained with a reasonable expenditure of money by any other means at present in use. Besides being of value in correlation work, electric logs have been found, in some instances, to be useful in estimating the nature of fluid contained within the pores of a rock. In special cases, even the relative permeabilities at different horizons have been determined.6 That electric logs should reflect in some measure such physical properties of rocks as porosity and permeability and, in addition, give rise to data pertaining to the fluid content of rock interstices, becomes evident when the conditions giving rise to the electrical phenomena measured are known. However, it is not the purpose of this paper to review the present status of electric logging; rather, the reader is referred to the papers cited on page 220 for the principles and practice involved. Importance of Drilling FLuid in Electric Logging Because the drilling fluid must be the medium through which the electric logging device is made to traverse the formations penetrated by a borehole, and since, in general, the drilling mud completely surrounds

Jan 1, 1942

By E. J. Duggan

The mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. Description of Ore The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chal-copyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. History Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From r926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for inolybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12.500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. TO increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. Crushing There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. a).

Jan 1, 1943

By Karl Fetters, John Chipman

The relationship between the composition of the slag and that of the underlying metal during the refining of a heat of liquid steel may best be studied in the light of the two broad physicochemical concepts of equilibrium and reaction rate. Actual steel-making processes, typified by basic open-hearth practice, involve such complexities of operating variables that it is impossible to distinguish with any degree of certainty between the general limiting conditions imposed by chemical equilibrium and the effects associated with rates of diffusion or of chemical reaction. Further advances in the physical chemistry of steelmaking will depend upon our ability to distinguish clearly between these two kinds of effects, and the experimental work that must be done falls naturally into two classifications. The first is the study of chemical equilibrium that cannot be carried out with assurance under operating conditions but must be done in the laboratory. The second is the -subject of rates of reaction and diffusion and the mechanism of the process as a whole, which can best be studied under operating conditions. The latter is not entirely independent of the former, since its interpretation often requires a knowledge of the equilibrium state. It is the purpose of this and of several papers that will follow to present the results of labora- tory investigations of equilibria involving liquid steel and slag. Open-hearth slag is a complex mixture containing at least eight major components and perhaps twice as many whose influence may be classified as probably minor. It is not surprising that progress toward a complete understanding of such a material at high temperatures has been slow. Attempts to apply the established principles of physical chemistry to the behavior of the individual components of this mixture have been beset by an almost complete lack of data on its chemical properties in the liquid state. This has necessitated the adoption of various methods of approximation based upon such assumptions as, for example, that for a given component the ratio of its mol fraction in the slag to that in the metal is a constant, or that the molecular composition of the liquid slag is the same as its crystalline constitution after solidification. From the thermo-dynamic viewpoint, such assumptions are merely useful dodges, which enable one to hazard a guess as to the magnitude of the more fundamental quantities, the chemical potentials and activities of the components of slag. The problem of determining the activity of each substance in an eight-component svstem seems too complex to be undertaken without the aid of generalizations derived from experimental data on similar systems containing a smaller number of components. Accordingly, it was planned in the present investigation to study the equilibrium of molten iron with the simplest

Jan 1, 1941

By W. A. Johnson, F. N. Rhines, W. A. Anderson

The rate of the high-temperature oxidation of pure copper has been measured repeatedly by numerous investigators. It appears to be generally agreed: (I) that at constant temperature, after the initial stages of oxidation have been passed and the scale has attained a substantial depth, the film thickness increases in proportion to the square root of the time of oxidation; and (2) that for any given period of oxidation the logarithm of the square of the film thickness is proportional to the reciprocal of the absolute temperature of oxidation.' These relationships are explained upon the basis that the rate of diffusion of copper to the external surface of the film is the controlling factor in establishing the rate of growth. Extensive studies have been devoted also to the mechanism and rates of the early stages of oxidation. There is thus available a reasonably well established theory of the oxidation of pure copper upon which to construct an explanation of the more complex oxidation of copper alloys. Few general principles have as yet emerged from the chaos of observations that have been reported in the field of the oxidation of alloys.2 It is evident, however, that the copper alloys may be divided into two main groups, whose oxidation characteristics are distinguish- able. The first group, composed of the alloys with the noble metals, forms substantially pure copper oxide scales; the alloying element is concentrated within the residual metal, and sometimes particles of almost pure noble metal are found in the scales completely surrounded by copper oxide. In the second group, composed of the alloys of copper with the baser metals, the alloying element appears as an oxide in the scale, often in greater concentration than in the alloy itself and sometimes to the exclusion of copper oxide. The oxide scales of the latter class have been found to be made up of one, two, and sometimes more distinguishable layers of which the innermost is often metallic in character, composed of particles of the oxide of the alloying element embedded in essentially pure copper and known as the subscale or zone of internal oxidation. The need for a thorough quantitative study of these modes of oxidation is evident. Since it is apparent that several mechanisms must be cooperating to produce the complex oxide scales found on copper alloys, the analysis of the process into its several simple operations should prove instructive. Fortunately, it is possible experimentally to produce internal oxidation in the absence of an external scale; this permits the isolation and intensive study of one of the steps in the oxidation process. In the present research the rates of internal oxidation of an extensive series of dilute solid solution alloys of copper have been measured and a theory has been developed to predict the observed

Jan 1, 1942

By A. L. Marshall, F. J Norton

The object of this investigation was to make a comprehensive study of the degassing of molybdenum in order to determine how rigorous a treatment was necessary to completely remove sorbed gases from molybdenum electrodes in vacuum tubes. Some work has also been done with tungsten, nickel, iron and carbon. As a logical development of this work it has been shown that the gas is present throughout the body of the metal and a study has been made of the solubility of nitrogen in molybdenum and tungsten and the rate, of diffusion of nitrogen through molybdenum has been calculated. Tests were made on molybdenum from all the commercial sources and no difference was discovered in the nature of the gases present, the amounts or the ease of removal. The reason for this was apparent when a study was made of the manner in which gases are absorbed by gas-free molybdenum. It is necessary to heat molybdenum to 1760°C. in a vacuum of the order of 0.001 micron for a time which varies lineally with the thickness, in order to obtain a condition in which no further gas is evolved by the sample. A sample so degassed can be handled and subsequently assembled in a tube and then degassed readily. The gases obtained from molybdenum are carbon monoxide and nitrogen, the nitrogen being the more difficult to remove. Apparatus and Technique The apparatus is shown in Figs. I and 2. It consisted of two units, a degassing system and an analytical system. Each had a bake-out oven, as shown surrounding the systems in the sketches. The system , will be traced, starting from the sample in the degassing system. A water-cooled quartz chamber surrounded the sample. The high-frequency coil, which heated the sample, extended only around this chamber and was cooled by an air blast. High-frequency power of 1800-meter wave length was obtained from a self-rectifying oscillator circuit with a maximum power input of 20 kw. A graded seal connected the quartz to the rest of the apparatus, which was of hard glass. The temperature of the molybdenum sample was observed by an optical pyrometer through an optical window about 40 cm. distant from the sample. Since the object viewed under these conditions was not a black body, true temperature was not observed, but the so-called "brightness temperature." The correction of this to true temperature is discussed in a later section. The apparatus was taken apart just above the graded seal; the quartz chamber was slipped off, washed with nitric acid, water and alcohol, and dried by an air stream. The old sample was removed by unscrewing the lower nut, and the new

Jan 1, 1944

By Robert S. Moehlman

Is recent years, exploration and development by diamond drilling has become a popular attack on the problem of finding new ore shoots and developing extensions of known shoots. Diamond drilling has been used in recent years in extensive sampling of disseminated copper deposits, in preference to churn drilling. At the Harvard gold mine, near Jamestown, California, Philip R. Bradley, Jr., general superintendent, reports that: "The Harvard work was a singular example of the reliance that can be placed on diamond drilling in evaluating an ore body." Drilling at the Harvard mine will be discussed in a subsequent section of the paper. So popular has diamond drilling become, indeed, that many who have had slight experience with drilling are inclined to regard it as a cheap cure-all. This paper proposes to show some of the limitations on the application of drilling and some of the steps that should be followed in drilling programs to obtain reliable results. Cost oF Drilling Drilling has the advantage, over shaft sinking and drifting, of lower cost and greater speed. As compared with churn drilling, the advantage of bringing up a coherent sample instead of fine cuttings for a sample is obvious. Diamond drilling is generally more expensive, however, and more engineering effort is required in combining the average of core and sludge samples. Exploration of drift walls by hammer drills to a depth of less than 30 ft. is generally cheaper than diamond drilling, but beyond this depth the diamond drill is cheaper and more efficient. The principal factor in determining cost of drilling is the character of the ground drilled. Broken ground with many fissures is most expensive to drill. Hard ground is less of a problem with diamond drills than with ordinary machine drills. Hardness or softness affects only the speed of drilling. Caving, easily washed, sandy or clayey ground requires continual cementing and casing, which is the greatest cost problem. Thus, soft, incompetent ground is more costly to drill than hard, massive ground. Cost of drilling per foot in good to fair ground varies from less than $I per foot, achieved in some districts in Ontario, the Mother Lode, and Missouri, to $5 per foot in many western states. In bad drilling ground, costs may range from $5 to $15 or more per foot, thus approaching the cost of drifting. Other factors that increase the cost per foot are: I. Short total length of hole drilled, so that the cost of moving in crews and equipment must be spread over a relatively few feet. 2. Cost of drilling per foot increases with depth beyond about 500 ft.. as more time is consumed in pulling rods' 3. Cost of preparing drilling stations, getting air underground to stations? etc' 4. Inaccessibility and lack of water on the job make for higher costs.

Jan 1, 1946

By R. S. Moehlman

IN recent years, exploration and development by diamond drilling has become a popular attack on the problem of finding new ore shoots and developing extensions of known shoots. Diamond drilling has been used in recent years in extensive sampling of disseminated copper deposits, in preference to churn drilling. At the Harvard gold mine, near Jamestown, California, Philip R. Bradley, Jr., general superintendent, reports that: "The Harvard work was a singular example of the reliance that can be placed on diamond drilling in evaluating an ore body." Drilling at the Harvard mine will be discussed in a subsequent section of the paper. So popular has diamond drilling become, indeed, that many who have had slight experience with drilling are inclined to regard it as a cheap cure-all. This paper proposes to show some of the limitations on the application of drilling and some of the steps that should be followed in drilling programs to obtain reliable results. COST OF DRILLING Drilling has the advantage, over shaft sinking and drifting, of lower cost and greater speed. As compared with churn drilling, the advantage of bringing up a coherent sample instead of fine cuttings for a sample is obvious. Diamond drilling is generally more expensive, however, and more engineering effort is required in combining the average of core and sludge samples. Exploration of drift walls by hammer drills to a depth of less than 50 ft. is generally cheaper than diamond drilling, but beyond this depth the diamond drill is cheaper and more efficient. The principal factor in determining cost of drilling is the character of the ground drilled. Broken ground with many fissures is most expensive to drill. Hard ground is less of a problem with diamond drills than with ordinary machine drills. Hardness or softness affects only the speed of drilling. Caving, easily washed, sandy or clayey ground requires continual cementing and casing, which is the greatest cost problem. Thus, soft, incompetent ground is more costly to drill than hard, massive ground. Cost of drilling per foot in good to fair ground varies from less than $1 per foot, achieved in some districts in Ontario, the Mother Lode, and Missouri, to $5 per foot in many western states. In bad drilling ground, costs may range from $5 to $15 or more per foot, thus approaching the cost of drifting. Other factors that increase the cost per foot are: I. Short total length of hole drilled, so that the cost of moving in crews and equipment must be spread over a relatively few feet. 2. Cost of drilling per foot increases with depth beyond about 500 ft., as more time is consumed in pulling rods. 3. Cost of preparing drilling stations, getting air underground to stations, etc. 4. Inaccessibility and lack of water on the job make for higher costs.

Jan 1, 1946

By R. Casting, J. Philibert, and C. Crussard

APPARATUS described in this paper uses the properties of X-radiation, emitted by substances under electron bombardment, as a means of rapid chemical point analysis. The method is based upon the principles of emission X-ray spectrography pioneered by Moseley in 1913. A very finely focused beam of electrons—the electron probe—is projected on the surface of a sample, which has a very large area compared with the beam diameter, at the point where it is desired to know the chemical composition. The minute volume of the sample which is thus irradiated emits a complex X-ray spectrum consisting principally of the characteristic radiations of the elements present in the volume. Measurement of the wavelength and intensity of each component of this spectrum thus affords a simple method of determining the chemical identity and concentration of these elements. The volume of sample which is analyzed at each position of the beam is limited by the diffusion of electrons in the sample. For an accelerating potential of the electron beam of about 30 kv, which is a convenient value for producing the characteristic radiation, this diffusion extends over a diameter of about 1 µ normal to the beam, and about 2 µ in the direction of the beam. Thus a beam of 1 µ in diameter leads to a volume of about 2 µ in diameter being irradiated. The limits of the analyzed region are thus considerably smaller than those obtained by using spark or arc emission spectrometry, which must extend to about 50 µ at right angles to the beam, and those obtained in X-ray fluorescence analysis, where the penetration of the specimen is considerably greater. The apparatus itself consists of three main parts: an electron optical system. an X-ray spectrograph and an optical microscope. The electron opticai system consists of a tungsten cathode emitting the electrow and a pair of magnetic lenses which focus the electron beam on the specimen to be analyzed. The specimen is prepared in the usual way, as in optical micrography. The X-radiation, emitted at the focal spot where the electrons strike the sample, is analyzed with a curved crystal, Geiger-Müller counter, vacuum spectrograph. The light microscope, designed to observe the surface of the sample in the region to be analyzed, is provided with a reflecting objective placed on the axis of the second magnetic lens. Its magnification is X450. The quantitative determination of an element A in the analyzed region is made by measuring the intensity I, of a strong characteristic line of the element, as emitted from the sample, and then the intensity I (A) of this same line as emitted from a

Jan 1, 1958

By R. T. Webber, R. B. Owen

The vapor phase deposition of molybdenum and tungsten from the hexacar-bonyls has been thoroughly investigated by Lander and Germerl, and shown to yield well-bonded coatings on a variety of hot metallic surfaces. They also demonstrated that chromium coatings2'3 could be similarly deposited, but did not investigate this metal extensively. Recently, conditions have been described by which this compound can be prepared in quantities sufficient for laboratory scale plating experiments, and it is the purpose of the present communication to outline the results of preliminary experiments upon the plating of chromium from its carbonyl. Apparatus and Technique The apparatus used in these experiments is shown in Fig I. At one end, connected by a short length of rubber tubing, is a Cenco Hypervac 20 pumping continuously. At the opposite end is a needle valve which allows a small stream of purified hydrogen to flow through the apparatus. Connected in series are a trap T1, maintained at the temperature of dry ice and acetone, the plating chamber, and two traps T² and T3 containing crystals of the carbonyl. T² is kept in a constant temperature bath (not shown), while T³ is at room temperature. Stopcock S6 allows the hydrogen flow to by-pass the carbonyl traps. A McLeod gauge is connected so that it can measure the pressure at the pump through Se, or the pressure at the plating chamber through S9 and S6. The standard (12 liter) volume connected by Sb is used to determine the volume included by S4, S6, SB, Slot the needle valve and the McLeod gauge. Knowledge of this volume allows calibration of the flow of hydrogen. The apparatus* is built entirely of pyrex, except for the needle valve, the rubber connection to the Hypervac, and the cover of the plating chamber. The plating chamber is water-jacketed, with a large brass top allowing insertion of the specimen to be plated. Set into the brass top is a ground glass stopper through which thermocouple leads are sealed. The thermocouple wires (pure Pt, and 87 pct Pt-13 pct Rh) serve the double purpose of measuring the specimen temperature and suspending it in the chamber. Surrounding the chamber is the coil of an induction heater, H. The apparatus is designed to allow variation of the following conditions: (I) the specimen temperature, from room temperature to about 1000°C, (2) the carbonyl vapor pressure, from 40 to 220 microns, and (3) the hydrogen flow. In several runs, 0.8 to 1.8 PCt of hydrogen sulfide was introuced through a needle valve that replaced S4. The vapor pressure of the carbonyl is

Jan 1, 1949

By Donald B. Hunter

The Ti-Be system up lo 10 wt pct Be with cortlmercialll' pure titanium has been determined using metal-lographic techniques. Beryllium forms the p eutec-toid type of dinqarz with titanium; the eutectoid occurs at 0.5 xt pct Be azd 1500 F, while an eutectic is foritled at 1887°F containing between 5 and 7.5 rt pct Be. The solubility of beryllium in a titanium does not exceed 0.05 zvt pct and the maximum solubility of Beryllim in 0 titanium is 1 wt pet. PREVIOUS work on the Ti-Be phase diagram indicated that additions of beryllium to titanium resulted in a marked lowering of the melting point,''' and it was also observed that the solubility of beryllium in both a and p titanium was retricted.- With the limited data available, McQuillan and McQuillan constructed a partial hypothetical diagram of the Ti-Be system,= assuming that beryllium formed the 0 eutectoid type of system with titanium. Subsequently Bedford published a tentative diagram of the entire Ti-Be system: although he did not investigate alloys containing less than 10 wt pct Be. This diagram included several compounds described by other authorities.'-l3 The following account describes work directed toward determining the form of the titanium-rich portion of the phase diagram up to 10 wt pct Be. MATERIALS AND METHODS Titanium sponge and a master alloy containing 25 wt pct Be were used in the formulation of the alloys; the chemistry of these materials is detailed in Table I. The alloys were prepared as 250-g compacts which were melted into buttons by arc melting under purified argon on a water-cooled copper hearth. Results of analyses for beryllium are shown in Table I but no determination of other elements was made at this stage. The alloys were hot-rolled at 1650°F to produce 0.1-in.-thick sheet. Alloys containing up to 1.5 pct Be were fabricated without trouble; samples containing 2 pct cracked initially upon hot rolling, but thereafter rolled satisfactorily to sheet. The 5, 7.5, and 10 pct were hot short and could not be rolled. Surface contamination was removed from samples by sandblasting and pickling in a 5 pct HF-35 pct HNOs solution. For metal log raphic examination, heat-treated specimens were mounted transverse to the rolling direction, ground on silicon carbide papers of increasing fineness to 600 grit, and electropolished at 20 v using a solution containing 600 ml of methanol, 60 ml of perchloric acid, 360 ml of butyl cellosolve, and 2 ml of Solvent "X". Difficulty was experienced in obtaining good electropolished surfaces with samples containing more than 1 pct Be, the degree of difficulty in-

Jan 1, 1967