Search Documents

For the year ending February, 1918 PRESIDENT PHILIP N. MOORE, ST. Louis, Mo. PAST PRESIDENTS WILLIAM L. SAUNDERS NEW YORK, N. Y. L. D. RICKETTS, NEW YORK, N. Y. FIRST VICE-PRESIDENT SIDNEY J. JENNINGS, NEW YORK, N. Y. TREASURER GEORGE C. STONE, NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND, NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON ` . . . . NEW YORK, N. Y.

Jan 7, 1917

By K. BRUCE DAVIDSON

-The North Central Texas area includes the following counties in Railroad Commission of Texas District 7-B: Brown, Callahan, Coleman, Comanche, Coryell, Eastland, Erath, Fisher, Hamilton, Haskell, Hood, Jones, Lampasas, Mills, Nolan, Palo Pinto, Parker, S

Jan 1, 1946

By H. B. Fuqua, B. E. Thompson

The area discussed herein, commonly known as the North Texas district, embraces the following 10 counties: Archer, Baylor, Clay, Cooke, Foard, Hardeman, Knox, Montague, Wichita and Wilbarger. It is underlain by two major structural features. The northernmost portion of the district, including Cooke, Montague, northern Clay, Wichita, Wilbarger, Foard and Hardeman counties, is underlain by the system of buried mountains known as the Red River uplift, which is parallel to and closely related to the Wichita-Amarillo uplift of Oklahoma and the Texas Panhandle. While the presence of a great majority of the numerous oil fields of the area is traceable to this feature, the production in Archer, southwestern Clay and southeastern Baylor counties is due to the influence of what is known as the Cisco Arch (so called because of its influence on Cisco deposition), the axis of which extends in a northwest-southeast direction across Archer County. It has been impossible for the authors to isolate the statistics for production for each of the many small pools of this district, as such statistics are not kept by the operators except for the district as a whole, therefore Tables 1 and 2 are comparatively short. The number of wells

Jan 1, 1935

By P. G. Winchell, M. Cohen, J. Woodilla

The kinetics of stabilization have been studied with respect to the isothermal component of the martensitic reaction in ivon-nickel alloys. Although the carbun (or nit-vogen) content may be very low in these alloys, it plays an important vole in the stabilization phenomenon. In fact, if these interstitial elements are removed by moist-hydrogen treatment, no evidence of stabilization is found. In the pvesence of 0.007 wt pct C, the activation energy for the stabilization process is comparable to that -for the diffusion of carbon (or nitrogen) in ferrite) rather than in austenitc. This suggests that the interstitial diffiision controlling the stabilization occurs within the martensitic embryos rather than in the matrix of the parent phase. The kinetics furthev indicate that the interstitial atoms tend to diffuse -from the embryos toward the suvroundings, thereby immobilizing the interface. This appears to be the origin of the stabilizing effect -found in these alloys. PHILIBERT' has shown recently that interstitial elements play an important role in the stabilization of martensitic transformations in iron-base alloys. However, determination of the temperature dependence for this stabilization process has produced values (- 11,000 cal per mol) that cannot be readily identified with the activation energy for diffusion of carbon (or nitrogen) in either ferrite or austenite. In the present work, the isothermal mode of the martensitic reaction was employed to derive the activation energy for stabilization, and this was found to agree fairly well with that for diffusion of carbon (or nitrogen) in ferrite. An iron-nickel alloy with 30.8 wt pct Ni and 0.007 wt pct C was used in these experiments, the specimens being in the form of rods 0.074 in. in diam and 2 1/2 in. long. The specimens were sealed in evacuated vycor tubes and austenitized at 1100°C for 1/2 hr prior to quenching in water. A second series of specimens was decarburized (and denitrogenized) prior to the above austenitizing treatment by annealing in moist hydrogen for 182 hr at 1215°C. In order to achieve a "standard" condition in the austenitic state, all the specimens were transformed to martensite by cooling to - 195°C before the final austenitizing. Both the 0.007 pct C and the decarburized series then had an Mi temperature of about - 30°C. The isothermal martensitic transformation was carried out at a reference temperature of -88"C, and was traced by means of electrical resistance measurements. Each resistance value was normalized by dividing by the resistance of the austenitic specimen at O°C, thus giving a resist- ance ratio R. The rate of decrease of the resistance ratio (-dR/dt) provided a measure of the transformation rate. It was found that a plot of R vs log t at the reference temperature resulted in a slightly curved line which could be closely approximated by two straight lines intersecting at a time of 10 min, as shown in Fig. 1. If G = dR/dlogt is the linear slope of either straight segment, the rate of transformation at any given time is: The ratio of G for 1 less than 10 min to that for t greater than 10 min was 1.19 for the 0.007 pct C series and 1.24 for the decarburized series. These ratios made it possible to compute the transformation rate at any time greater than 10 min, knowing the transformation kinetics prior to 10 min. The stabilization treatment was introduced by up-quenching from the reference temperature, usually after 10 min of isothermal transformation: to one of four stabilizing temperatures between 0" and 22"C. After holding for various times, the specimens were down-quenched to the reference temperature for additional isothermal transformation. In this way, it was possible to compare the transformation rate (- dR/dt), immediately after stabilization with the transformation rate (- dR/dt)U in the unstabilized condition. The degree of stabilization was defined as: Ordinarily, Gs and Gu would be measured after the same time of prim isothermal transformation.

Jan 1, 1960

By Ivan E. Goodner, Charles H. Fulton

THE subject of the constitution of copper-iron mattes has received considerable attention in recent years by Keller,1 Bolles,2 Hofman,3 and Gibb and Philp.4 Still more recently Friedrich, Röntgen, Wiedmann, Heyn and Bauer, and Leroux have published, in some instances, elaborate researches on the various systems of the sulphides, and the sulphides and their respective metals, which enter into the constitution of mattes, and have added very greatly to the knowledge of the subject. The work of Friedrich has been especially fruitful. But to date no effort has been made to correlate the data obtained in the latter researches and apply then to the study of the product called " matte," which is made in such large quantities by copper and lead-smelters. In this paper we propose to determine, by metallographic and chemical means, the constitution of copper-iron and copper-lead mattes in the light of the recent data above mentioned. We do not discuss copper-nickel mattes or speisses (artificial arsenides). The paper is divided into three sections I. A review and discussion of the various binary systems of the metallic, sulphides, and the binary systems of various metals and their respective sulphides entering into the constitution of mattes; also a discussion of the physical and chemical properties of these sulphides. II. A description of the results obtained from the chemical and metallographic examinations of the samples of copper-iron and copper-lead-iron matte obtained from various plants.

Nov 1, 1908

For the year ending February, 1918 PRESIDENT PHILIP N. MOORE, ST. LOUIS, Mo. PAST PRESIDENTS WILLIAM L. SAUNDERS, NEW YORK, N. Y. L. D. RICKETTS, NEW YORK, N. Y. FIRST VICE-PRESIDENT SIDNEY J. JENNINGS NEW YORK, N. Y. TREASURER GEORGE C. STONE, NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND, NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON NEW YORK, N. Y.

Jan 2, 1918

For the year ending February, 1918 PRESIDENT PHILIP N. MOORE, ST. Louis, Mo. PAST 'PRESIDENTS WILLIAM L. SAUNDERS NEW YORK,. N. Y. L. D. RICKETTS NEW YORK, N. Y. FIRST VICE-PRESIDENT SIDNEY J. JENNINGS NEW YORK, N. Y. TREASURER GEORGE C. STONE, NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND, _ NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON NEW YORK, N. Y.

Jan 11, 1917

By K. Sallmann

Spurred by a variety of factors, coal flotation is making headway among the preparation plants of West Germany. The author gives some statistics on German coal flotation plants and information on the properties of the feed and quality of the derived products. Types of machines and reagents used, dewatering practices, tailings disposal, and thickening operations are covered. There is special emphasis on the many-sided problems that confront German coal preparation engineers. The coal mining industry of West Germany is concentrated in three coalfields: the first on the Ruhr and the Rhein rivers, the second on the Saar river, and the third near the frontier between Germany and the Netherlands, around the towns of Aachen and Erkelenz. The total run-of-mine production of these three coalfields amounts to 736,000 tpd which come from 145 collieries. There are 121 washeries, 43 of them with a flotation plant. The total throughput of these 43 washeries is 280,000 tpd, out of which 26,000 tons (9.3 pet of the washery feed coal) are cleaned by flotation. The capacity of the individual flotation plants varies within relatively wide limits, the average being 40 tph and the capacity of the largest flotation plant being 120 tph. During the last few years, the number and size of flotation plants have been steadily increasing, although flotation must be looked upon as an expensive and rather complex cleaning process. The considerations which have led to the widespread application of flotation may be summarized as follows: 1) In making coking coal, it is seldom possible to add uncleaned fines to the coke oven charge, as their ash content is too high. If, however, the ash content of the fines is reduced to a maximum of 7 or 8 pet, they can be admixed to the washed and crushed small sizes without difficulty. This means, naturally that revenue is increased since the price of coking fines is always higher than the price for dust or uncleaned wet fines. 2) To prevent silicosis and pneumoconiosis, water infusion and spraying of the coal are practiced much more today than ever before. As a result, moisture content of run-of-mine coal has markedly increased and the washery feed coal contains more and more slurry instead of dry dust. The price of slurry, how- ever, is very low, and in many cases it is impossible to sell filtered raw slurry with a moisture content of 20 to 25 pct and, at the same time, 20 to 25 pct of ash. Reduction of the ash content of these slurries improves the possibility of dewatering them and, in this way, also enhances the marketability of this product. 3) The new severe laws and regulations against pollution of air and rivers make it necessary to de-dust coal better than in the past and to reduce the quantity of the waste water from coal preparation plants and their solids content to a minimum. 4) Even if there were no compulsory reasons to clean the fines, flotation will often lead to an increase in the overall yield. Although not the only application, it appears that the treatment of coking fines is the primary field for the flotation process in coal preparation. It must not be overlooked, however, that a series of arguments may be advanced against flotation: 1) Flotation is an expensive process because, in addition to the cleaning operation itself, the dewatering of the froth and the disposal of the tailings is very costly. 2) Operation of a flotation plant requires well trained personnel. 3) Filtered froth has about a 20 pct moisture content and, therefore, if mixed with the cleaned small coal it causes an undesirable increase in water content of the coke oven charge. For this reason it is necessary to take additional measures for dewatering washed small coal and this, of course, entails additional expense. 4) Disposal of flotation tailings involves very difficult problems, particularly in the highly industrialized regions. An investigation covering 32 flotation plants of Western Germany shows the following composition and properties of their feed:

Jan 1, 1961

By J. P. Boisseau, C. N. J. Wagner, E. N. Aqua

An analysis ulas made of powder-paltern peaks from cold-rolled polycrystalline copper and from copper powders, compacted into briquets (1 in. diameter) with pressures up to about 106 psi. Powder-pattern peaks from both contact faces were measured with a diffractometer and CuKa radiation. Laltice-parameter measurements indicate the presence of residual stresses in the bulk specimen of the order of a - -10 kg per sq mm after 93 pct reduction in thickness, and give some evidence for the occurrence of deformation stacking faults (probability a - 0.003) in both bulk and powder samples. Fourier analyses of the line profiles were carried out on the (111) - (222) and (200) - (400) Pairs of reflections. The particle sizes decrease rapidly with increasing deformation and reach asymptotic values of De (111) = 550A and D, (200) = 350A at about 25 pct reduction in thickness or 80,000 psi pressure for compaction. The anisotropy of the particle sizes indicates the presence of microtwins and stacking faults on (111) planes. The rms strains increase also with deformation and reach limiting values of 2 x 10-3 in the [100] direction and 1 x 10-3 in the [Ill] direction at a distance L = 100A normal to the reflecting planes. THE diffraction of X-rays in cold-worked, polycrystalline, fcc metals and alloys yields Debye-Scherrer patterns in which the individual (hkl) lines may show a broadening, a shift, and an asymmetric shape.' These effects can be explained by the fragmentation of the small, coherently diffracting domains, the presence of residual elastic strains in the lattice, and the formation of stacking faults. X-ray studies of stacking faults in fcc metals and alloys have been methodically performed during the last few years, but so far most of the investigators dealt with cold-worked powder specimens obtained by filing.2-8,10-19 In such specimens, the crystals are randomly oriented and the average elastic strains are assumed to be zero, so that the observed peak shifts are due to stacking faults only. The agreement between X-ray diffraction theory1,20 and experiment is generally good. On the contrary, cold-worked, polycrystalline bulk specimens (after wire drawing, rolling, plastic deformation by tension or compression, etc.) contain residual elastic strains, which may not average to zero in the surface of the specimen and show very strong preferred orientation of certain lattice planes (texture) parallel to the specimen surface. Nonzero elastic strains also produce peak shifts, the magnitudes of which depend upon the crystallographic orientation of the reflecting lattice planes in elastically anisotropic materials as well as on the type of plastic deformation and which have to be taken into account when measuring stacking-fault probabilities in deformed bulk specimens.21-23 The purpose of the present investigations was to perform a detailed X-ray analysis on a bulk specimen of a fcc metal deformed by rolling at room temperature and on powder of the same material deformed by compaction, and to compare these results with those previously obtained on cold-worked filings.18 Pure copper was chosen for this study since it is known to produce only a small quantity of stacking faults by filing at room temperature.3,16,17,19 The residual elastic strain in the surface of the bulk copper should be the predominant factor determining

Jan 1, 1965

For the year ending February, 1918 PRESIDENT PHILIP N. MOORE, ST. Louis, Mo. PAST PRESIDENTS WILLIAM L. SAUNDERS ? NEW YORK, .N: Y L. D. RICKETTS NEW YORK, N. Y. FIRST VICE-PRESIDENT SIDNEY J. JENNINGS . . NEW YORK, N. Y. TREASURER GEORGE C: STONE, NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND, . . . NEW YORK, N. Y. SECRETARY' BRADLEY STOUGHTON NEW YORK, N. Y.

Jan 6, 1917

By Heinrich Ries

I. INTRODUCTION. THE facts is presented in this paper, based chiefly on recon¬naissance made, during the summer of 1903, by myself and my assistant, Mr. R.. C. Brooks, cover practically all that portion of Texas lying east of the 99th meridian. The work was undertaken for the University of Texas Mineral Survey, and was to have been extended over the entire State, but the sudden termination of the survey prevented even the official publication of the results already obtained. Most of the mineral resources of Texas were fully treated by the First Geological Survey, but the clay-industry was tit that time little developed, so that only a few scattered notes concerning it are to be found in the Survey reports. The extensive exploitation of the clay-deposits, however, has emphasized their commercial importance, and has also made accessible many facts of exceptional interest concerning their geologic conditions. The erection of new plants for the utilization of clay-products is usually preceded by more or less prospecting, which, by reason of the scarcity of outcrops, and the geologic and topographic conditions of eastern and southeastern Texas, is often slow and difficult work. This is especially true in the Tertiary and Pleistocene areas, where the structural conditions much resemble those of the Atlantic coastal plain, the deposits being mostly lenticular in form, and surrounded by beds of sand. II. GENERAL GEOLOGY OF THE CLAY-DEPOSITS. The accompanying sketch-map, Fig. 1, shows the location of nearly all the deposits examined, their relation to the geology of the State, and the type of clay found in each. The geology is based on the published work of Hill, Cummins, Tail, Adams, Hayes and Kennedy, but the boundaries have unfortunately not been determined over the entire region covered by the map, and, in many areas, are only approximate. It will be seen that the clay-deposits range from Carboniferous to Pleistocene in age, the older deposits being found in the northwestern part of the area, while those of Cretaceous and Tertiary age lie to the east, southeast and south. The Pleistocene clays are found in part in a belt along the coast, and in part along many of the larger rivers, where they often underlie extensive terraces.

Sep 1, 1906

By J. Washburn, R. Drouard, E. R. Parker

Temperature dependence of the rate of recovery in zinc single crystals after a simple shear deformation at low temperature was investigated. Some tentative suggestions regarding the annealed and strain-hardened states of a crystal are discussed. RECOVERY may be defined as the gradual return of the mechanical and physical properties of strain-hardened metal to those characteristic of the annealed material; an increase in temperature increases the rate of recovery. The annealing process in strain-hardened polycrystalline metals is complicated by the inhomogeneity of strain which always exists in aggregates. Polygonization in bent regions of the crystals and growth of new almost strain-free grains starting at points of severe local distortion1-:' make it almost impossible to isolate and study the recovery process. Homogeneously strained single crystals, however, do not polygonize or re-crystallize and hence they can be used advantageously to study recovery. In such crystals strain hardening is completely removed by recovery alone. Since recovery is a process whereby certain lattice disturbances introduced by plastic flow are gradually reduced, a knowledge of the rate and temperature dependence of this process for various conditions of prestrain might be helpful in formulating a model of the strain-hardened state. For simplicity it seemed desirable to limit the type of prestrain to the simplest obtainable, i.e., simple shear strain. In the experiments to be described, recovery was studied by observing changes in the stress-strain curve of prestrained zinc single crystals held for various times at temperatures above that employed for straining. Single crystals were grown from the melt by a modified Bridgeman technique from Horse Head Special zinc 99.99 pct pure, and from spectrographically pure zinc 99.999 pct pure. They were grown as 1 in. diameter spheres and acid-machined' to the final specimen contour. The test section was a cylinder about 1/8 in. high and 3/4 in. in diameter. The conical sections adjacent to the test section were cemented into the grips so the load could be transmitted to the crystal as uniformly as possible. The specimens were oriented so that in testing the maximum shear stress was applied along one of the slip directions, [2110], in the (0001) plane. Details of the production and testing of such specimens have been presented.' Each test was carried out according to the following schedule: 1—The crystal was strained at — 50°C until it reached a maximum shear stress, ,,,. The strain rate was approximately 5 pct per min in all cases. 2—After straining, the crystal was unloaded before the temperature was changed. Unloading required about 3 min. 3—The temperature of the specimen was then increased from — 50°C to the temperature, T, of recovery. This change in temperature was completed in a time of less than 2 min. The specimen remained at temperature, T, for a time, t, which differed for the various specimens. 4—Thereafter the temperature was again reduced to — 50 °C in approximately 3 min. 5—While at —50°C, the stress-strain curve after recovery was obtained. 6—The specimen was then unloaded and annealed for 1 hr at 375 °C in a helium atmosphere to bring about complete recovery. Cooling to room temperature after anneal required 90 min. 7—The same crystal could be re-used for another test because the plastic properties after annealing closely duplicated those of the original crystal. The specimen was immersed during the test in a bath of methyl alcohol which, through a system of tubes, could be pumped through either of two heat exchangers to regulate the temperature; this was accomplished by circulating the liquid through coils immersed in a bath of acetone and dry ice for cooling or in a bath of warm water for heating. Test temperatures were thus maintained constant within ±1°C. The — 50°C temperature was low enough so that no measurable recovery occurred during unloading and reloading. The stress-strain curve continued after recovery along a path below, but approximately parallel to, the path of a curve obtained in an uninterrupted test. Fig. 1 shows some of the results from a specimen of 99.999 pct Zn. The amount of downward displacement of the curve due to recovery was a

Jan 1, 1954

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

By C. H. Curtis

HE valuable mineral content of the current feed -*- to the Miami concentrator is as follows: copper, 0.7 pct total; molybdenum, 0.01. Flotation of this ore yields a sulphide concentrate containing: chalco- cite, 44 pct; molybdenite, 0.5; pyrite, 50.0; insol, 5.5. A combination of potassium ethyl xanthate and pentasol amyl xanthate as collectors, and pine oil as frother, are used in this flotation. Rejection of pyrite is encouraged by holding the amount of collectors used to the minimum consistent with copper recovery and by operating at high alkalinity (equivalent to 0.35-0.40 lb CaO per ton solution of pH 11.0). The molybdenum recovery in the sulphide concentrates under the above flotation conditions is approximately 50 pct of that originally present in the ore. Taking into account the acid soluble molybdenum, indicated molybdenite recovery is 75 to 80 pct. The attempt to separate the molybdenite into an acceptable molybdenum product begins with the bulk sulphide flotation concentrate just described. This concentrate is composed of chalcocite, whose floatability has been promoted to the fullest extent possible for the sake of its recovery from the ore, together with the pyrite which has been activated along with the copper mineral. The problem is to deaden the copper and iron minerals, and to float the molybdenite. Obviously in the accomplishment of this end, conditioning and preparation of the pulp, prior to flotation, plays an all important role. The first step is thickening to 50 to 60 pct solids, with milk of lime added to the thickener feed to maintain an alkalinity of the pulp equivalent to a pH of 8.5 to 8.8 during its residence in the thickener. The purpose of the thickening is primarily to reduce the volume of pulp for subsequent treatment. However, the relatively prolonged retention of the pulp in the thickener at the desired alkalinity is known to have a favorable depressing effect upon pyrite. There is a limit for this alkalinity above which a depressing effect upon molybdenite occurs. The thickened pulp (alkalinity: 0.015 lb CaO per ton, pH 8.8), discharges into an agitator, retention time approximately 2 hr, to which additional lime is added to raise the alkalinity to 0.35 to 0.40 lb CaO per ton solution, pH 11.6. This additional lime is required for pyrite depression and can be tolerated without loss of molybdenite because of the limited time of contact in the conditioner tank. The pulp leaving the lime conditioner passes through two successive steaming tanks, which are mechanically agitated, and into which live steam is admitted directly into the pulp near the bottom of the tanks. The temperature of the pulp is maintained as near boiling as possible. The steaming time is approximately 4 hr. The pulp leaving the last steamer has an alkalinity of about 0.04 lb Cao per ton solution, pH 8.7. It is believed that oxidation of the copper and iron sulphides occurs during steaming, the resulting sulphates reacting the calcium hydroxide to calcium sulphate and thus reducing the alkalinity. Since the steamer-feed solution is already saturated with calcium sulphate, the calcium sulphate produced during steaming is precipitated. It is believed that this calcium sulphate is precipitated preferentially on copper and iron mineral surfaces thus decreasing their floatability. Aside from the "lime chemistry" during steaming, pine oil is displaced from the pulp and xanthate decomposed, which has a major effect upon the deadening of the copper and iron sulphides. Following steaming, the hot pulp is admitted to another conditioning tank wherein it is aerated, primarily for cooling, but incidentally for additional oxidation of the copper and iron sulphides. The resulting "deadened" pulp is then diluted to 20 pct solids, a specific collector for molybdenite, ordinary stove oil, is added and the separation of the molybdenite by flotation is undertaken at a pH of 8.5 to 8.8 in standard Miami air-flotation ma-chines. B-22 frother is used when necessary. A re-grind of the thickened rougher concentrates is made prior to the first cleaning operation chiefly for rejection of insoluble in subsequent flotation. The cleaner concentrate is then stepped up to 90 pct MoS, in an 8-cell Denver flotation machine No. 18. Sodium silicate is added to the cleaner circuit. Its effect is to flocculate molybdenite and stabilize the froth. In summary, it may be stated: 1. Separation of molybdenite into an acceptable product from sulphide copper concentrates by flotation involves preliminary preparation and conditioning of the pulp, which is of major importance. 2. This preparation and conditioning consists of several successive steps: (A) Thickening to 50 to 60 pct solids at controlled alkalinity to reduce volume of pulp and to contribute to depression of pyrite. (B) Agitation at high-pulp density for limited time with additional lime to provide for depression of pyrite. (C) Steaming at high-pulp density for decomposition of xanthate and xanthate surface films, evolution of pine oil, and oxidation of sulphide minerals other than molybdenite. The latter involves sulphating of lime with probable precipitation of calcium sulphate preferentially on copper and iron minerals. (D) Aeration at high-pulp density for cooling, and for further oxidation of copper and iron sulphide minerals. (E) Dilution of pulp to 20 pct solids and addition of specific collector for molybdenite, common stove oil. It is hardly necessary to point out that this rather drastic procedure for depression of previously activated copper and iron sulphide minerals, without at the same time depressing molybdenite, is possible due to the inherently high floatability and refractory nature of molybdenite. However, molybdenite is susceptible to depression by excessive lime which must therefore be limited to the amount consistent with satisfactory molybdenite recovery. The steaming procedure is being carried on at Miami Copper Co. under license agreement with Janney, Nokes, and Johnson, holders of letters patent on the process.

Jan 1, 1951

The Second Pan-American Scientific Congress will be held under the auspices of the United States Government in Washington, D. C., Dec. 27, 1915, to Jan. 8, 1916. The organization officers are John Barrett, LL. D., Director General of the Pan-American Union, Secretary General; Glen Levin Swiggett, Ph. D., Assistant Secretary General; the headquarters of the Congress are the Pan-American Union, Washington, D. C. The program of the Congress is divided into nine main Sections, which are given below with the name of the Chairman of each Section: 1. Anthropology, William H. Holmes, B. S., Smithsonian Institution, Washington. 2. Astronomy, Meteorology, and Seismology, Robert S. Woodward, Ph. D., Carnegie Institution, Washington, D. C. 3. Conservation of National Resources, Agriculture, Irrigation and Forestry, George M. Rommel, B. S., Bureau of Animal Industry, Department of Agriculture, Washington, D. C. 4. Education, P. P. Claxton, LL. D., Bureau of Education, Washington, D. C.

Jan 12, 1915

By R. M. Hudson

Equilibrium constants have been determined for the jeaction C (in steel) + H2O = CO + H2 as a function of carbon content (0.013 to 0.74 ujt pct) and temperature (1200° to 1800°F) by using a flow system in which the water-vapor content of various gas mixtures containing hydrogen, carbon monoxide, and nitrogen was systematically controlled. The equilibrium partial-pressure ratio K1 = PcoPH2/PH2O was obtained for two low-carbon steels and four Fe-C binaries from a series of runs made at each temperature. By comparison of Ki with the analogous ratio for the reaction C (graphite) + H2O = CO + H2, the activity of carbon in steel (ac) was comDuted for each temperature and carbon content studied. From these data values of the equilibrium constants for two other carburization-decarburiza-tion reactions of importance were estimated. The chemical compositions of the two steels and four Fe-C binaries used are given in Table I. Specimens of Steel E were cut from a 0.035-in.-thick commercially box-annealed and temper-rolled lot of a low-carbon low-metalloid steel. Steel F was a commercially hot-rolled rimmed steel that was cold-reduced at the Laboratory to a 0.059-in. thickness. Fe-C binaries A, B, C, and D were vacuum-melted and poured, hot-rolled, pickled, and cold-reduced at the Laboratory to furnish specimens 0.035 in. thick. Sheet specimens were cut from all lots to approximately 3/4 by 3 in. by gage and were vapor-degreased with trichlo-roethylene and weighed before heat treatment. Heat treatments were carried out in a hinged-top Hevi-Duty electric furnace with a Vycor-glass tube. When the furnace was at the desired

Jan 1, 1965

By Alan U. Seybolt

In examination of some cold-rolled copper and cupronickel that contained a little columbium, it was discovered that these alloys were unusually resistant to annealing. This effect was suficiently pronounced to warrant some investigation of the annealing characteristics of such compositions. Experimental Procedure The alloys were made from cathode copper, electrolytic nickel, and a special copper-columbium hardener alloy furnished by the Electro Metallurgical Corporation. This hardener alloy contained 10 per cent columbium, with tantalum, aluminum and iron as major impurities. The melting was carried out in a clay-graphite crucible in a high-frequency induction furnace! under a charcoal or electrode carbon cover. No deoxidizing additions were made. The 7-lb. copper heats were poured into a steel mold, hot-rolled to 1/4-in. thickness, and annealed for 56 hr. at 760°C.; the cupronickels were annealed I hr. at 815°C. The material was then cold-rolled to a 50 per cent reduction in thickness and annealing tests were carried out on this 1/8-in. hard-rolled strip. A little silver was added to copper-columbium alloys 402 and 403 to find out

Jan 1, 1942

By Alan U. Seybolt

In examination of some cold-rolled copper and cupronickel that contained a little columbium, it was discovered that these alloys were unusually resistant to annealing. This effect was suficiently pronounced to warrant some investigation of the annealing characteristics of such compositions. Experimental Procedure The alloys were made from cathode copper, electrolytic nickel, and a special copper-columbium hardener alloy furnished by the Electro Metallurgical Corporation. This hardener alloy contained 10 per cent columbium, with tantalum, aluminum and iron as major impurities. The melting was carried out in a clay-graphite crucible in a high-frequency induction furnace! under a charcoal or electrode carbon cover. No deoxidizing additions were made. The 7-lb. copper heats were poured into a steel mold, hot-rolled to 1/4-in. thickness, and annealed for 56 hr. at 760°C.; the cupronickels were annealed I hr. at 815°C. The material was then cold-rolled to a 50 per cent reduction in thickness and annealing tests were carried out on this 1/8-in. hard-rolled strip. A little silver was added to copper-columbium alloys 402 and 403 to find out

Jan 1, 1942

By E. M. Norris

Phosphate deposits are distributed widely over the earth's surface. Of the known areas of deposit, eight fields are of particular interest because of their vast reserves of high grade phosphatic material and the volume of their production that goes to the world's great consuming areas. These fields, in the order of their present economic importance, are situated respectively in Florida, French Morocco, Kola Peninsula of U.S.S.R., Tunisia and Algeria, Tennessee, Montana-Idaho-Wyoming-Utah, Ocean and Naru Islands in the Gilbert group, and in Egypt. The Rocky Mountain deposits have a comparatively low ranking in the phosphate trade of today, but this field contains one of the largest known reserves of phosphatic material, associated with which are recoverable trace minerals, which are of increasing importance in our national economy. The first discoveries of phosphate rock in the western states were recorded in Cache1 and Rich2 counties in Utah, respectively in 1889 and 1897. Mining operations were begun on these deposits near Montpelier, Idaho, in 1906. During the latter year geological studies of the phosphate beds were started by Weeks and Ferrier of the U. S. Geological Survey. These studies have been continued to the present day by a distinguished roster of the Survey's technicians. As a result of this systematic exploratory program, the Survey has published a series of bulletins, papers, and maps, which provide the mining public with a comprehensive study of the exposed

Jan 1, 1949

Reading Anthracite, like AIME, is celebrating its 100th anniversary. While Reading marks a century in the mining, preparation and marketing of domestic anthracite, AIME commemorates 100 years of parallel endeavor in promoting the economical production of minerals and metals and the welfare of those employed in the industries dedicated to this pursuit. Both Reading and AIME possess another common denominator, in that both were conceived in the eastern anthracite fields of Pennsylvania, where coal reigned supreme in the early days of mining.

Jan 1, 1971