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By R. W. Heckel, H. P. Cheskis, A. Lawley, M. R. Pinnel

THERE are discrepancies reported in the literature regarding the micro and macro stress-strain behavior of metals and alloys. In particular, Young's modulus and micro yield measurements often show significant variations depending upon the techniques employed.'-" The purpose of this note is to correlate measurements made with four techniques and to indicate where some of these discrepancies might arise. Allied use of these techniques can provide a complete and consistent characterization of micro and macro stress-strain behavior (over the strain range of 10-9 to 10--2 in. in.-') from a minimum number of samples. Techniques. The techniques employed for measuring stress-strain behavior involved the use of a) linear differential transformers, b) SR-4 strain gages, c) longitudinal resonance waves, and d) X-ray diffraction. Specimen geometries and preparations are summarized in Table I for each technique. Differential Transformer and Strain Measurements. An Instron universal testing machine, modified with an alignment device, was used to apply the load via parallel-jaw grips. The signal from the load cell was applied to the Y scale of a Moseley X-Y recorder (sensitivity of 10 lb load per in. chart). The elongation was monitored simultaneously with strain gages mounted on three sides of the rectangular specimen, and with two linear variable differential transformers (LVDT) mounted at 90 deg to one another on faces that also contained the strain gages, Fig. 1. The signal from one of the LVDT's was applied to the X scale of the recorder (sensitivity 40 µ in. elongation per in. chart) in order to trace a continuous load-elonga-tion curve. Concurrently, the strain on the same face (i.e., the face monitored by the LVDT) was recorded by the corresponding strain gage. A stress-strain curve was plotted from the load cell-strain gage data and was compared to the stress-strain curve (converted from load-elongation) plotted from the continuous load cell-LVDT trace. Identical Table I. Description of Specimens Used in the Present Study Material Dimerisions Preparation Condition 99.99 pct A1 tensile 318 by 3/16 by 3 in. as cast ingot; cut annealed specimen on diamond wheel 2024 A1 tensile 318 by 3/20 by 31/2 in. vacuum hot annealed specimen pressed sheets; cut on diamond wheel 99.99 pct Al 3116 by 3/16 by 2 in. cut from tensile prestrained dynamic modulus specimen after 10"* in. specimen test per in. 2024 A1 dynamic 3/20 by 3/20 by 2 in. vacuum hot annealed modulus specimen pressed sheets; cut on diamond wheel 2024 A1 X-ray gage section 112 by vacuum hot annealed specimens 1/4by1/20in. pressed sheets; cut on diamond wheel

Jan 1, 1970

By R. Shuttleworth, R. Smith

A Knudsen effusion tnethod I~as been used lo measure the vapor pressure of pure iron in the temperature range 1000° to 1500°C. Neutron-irradiated , natural iron was used and the Mn'~proclzdced by the (n, p) reaction removed by evaporation at high temperature. The results can be represented by the equation: where pM - 1.088 x 107 atm and T, = 20,890°K. A value of 100.2 kcal per g- atom was obtained for AH;. THE vapor pressure of pure iron in the temperature range 1000" to 1500°C has been measured. Radioactive iron and a Knudsen cell technique were used and the Mn' produced by an (n, p) reaction removed by evaporation at high temperature. From the variation of vapor pressure with temperature values of the enthalpy and entropy of solid iron have been found. Previous workers2 used a Lang-muir method and measurements were made over a smaller temperature range than that in this work. EXPERIMENTAL The apparatus used was similar to that of Pyle and Shuttleworth, and McLellan and Shuttle worth. However, because of the higher temperatures the Knudsen cell was machined from tungsten bar and the flanged furnace tube was beryllia. Heating was by passing a high current (70 amp at 7 v for 1500°K) through a strip of tantalum foil laid across the top of the cell and through a turn of molybdenum wire around the furnace tube. This method of heating ensured that the top of the cell was slightly hotter than the base and prevented condensation and re-evaporation from the orifice. The temperature was measured by means of a pt/Pt 13 pct Rh thermocouple spot-welded to the base of the crucible. To ensure accurate temperature measurement the iron was in the form of a thin disc (0.75 cm diam by 0.075 cm) which lay on the base of the cell. The high-purity iron was irradiated in the Harwell pile for a week at a neutron flux of 1.2 X 10" neutrons per sq cm per sec. The isotopes fe'~ and Fe" are produced by (n! y) reactions but only Fe", half-life 45.1 days, is a emitter; Mn which is also a y emitter, half-life 300 days, is produced simultaneously by an (n,p) reaction.' Immediately after irradiation of natural iron the ratio of ~e" to Mn y activity was 100:l. However, manganese is very much more volatile than iron and almost all the y activity of the condensate is due to Mn'~. Thus it was essential to eliminate all the manganeses before any measurement of the vapor pressure of iron was made and this was done by preliminary heating for 20 hr at 1400°C. For each determination the iron was dissolved from the target with acid and the y activity compared to a known weight of iron. By means of a y-ray spectrometer the y-activity measurements were made at the 1.09-mev peak of the 17e5' spectrum where the background was of the order of 3 cpm. Despite the low specific activity of ~e" a mass of 10- " g of iron could be determined corresponding to a vapor pressure of about l0-'' atm. RESULTS The vapor pressures were calculated from the equation: Where m is the mass of iron effusing in a time t through an orifice of area a2, r is the radius of the target, Z is the distance from the orifice to the

Jan 1, 1965

SAN FRANCISCO SECTION W. H. SHOCKLEY, Chairman, PRANK H. PROBERT, Vice-Chairman, T. A. RICHARD, Sec.-Treas., 420 Market St., San Francisco, Cal. C. C. BRAYTON, J. F. NEWSOM. The San Francisco Section met on September 11 to discuss the problem of Labor in its Relation to the Mining Engineer. Mr. W. H. Shockley was in the chair. T. A. Rickard opened the discussion under four heads: The recognition of the union, the rustling-card, the sliding scale of wages, and the' contract system. He expressed his belief that collective bargaining cannot be ignored, and insisted that the mining company represented a collective bargaining of its shareholders with the men. employed in its mine. He held that the contract system was often used unfairly, but that used fairly it was the best check on labor tyranny. The rustling-card he held to be right in principle, in so far as it gave the employer the choice of employees, but he believed that it was often used in an irritating way to black-list refractory laborers. The sliding scale of wages was easy to establish in prosperous times, but it would inevitably lead to trouble when conditions went against the miner. His conclusion was that the quarrels of labor and capital were due mainly to a lack of the sense of fair play and of social justice. Albert Burch agreed with most of the foregoing, and insisted that 90 per cent. of the men were inclined to be just, but that the 10 per cent: made all the trouble. He stated that he had introduced the rustling-card in the Coeur d'Alene in 1899, as a means for getting rid of agitators. E. L. Foucar considered that labor was irresponsible. He described his experience in Peru, where the authorities promptly stopped a strike of Italian miners by threatening prompt punishment for any display of violence. Morton Webber also gave a bit of his own experience, to show that the chief trouble is that the men do not think for themselves, and are therefore easily led by fomenters of trouble. F. H. Probert described his experiences in Germany and. also in Arizona, and insisted that too often the men were not given a fair hearing. E. T. Blake argued that a manager must get the workmen's point of view, in order to control them effectively. He believed that young men who started life doing manual labor were in a better position to understand their fellow workers. George J. Young suggested that the mine owners should encourage the good men to lead the unions, and not leave the direction of the unions to the noisy element.

Jan 12, 1917

By J. L. Gillson

In a brief summary of the many occurrences of fluorspar in our western states, it is not possible to go into detail in regard to the geology, mining and milling methods, and reserves about individual deposits. Rather a broad review is made of the occurrences and an analysis of how the western deposits may fit into our future domestic economy is presented. General Statement on Fluorspar Situation Fluorspar has been used for years as a flux in the manufacture of basic open-hearth steel, in foundry cupolas for melting pig iron, in the manufacture of opalescent glass and in the preparation of hydrofluoric acid. In 1880 the annual production of fluorspar in the United States was around 4000 tons per year. By 1900 it was up to 21,656 tons and between 1902 and 1908 it averaged about 40,000 tons annually. Prices were then so low, by present standards, that it seems impossible to believe now that mining could have been performed profitably at the price levels then prevailing. In 1890 fluorspar sold at the mines for $6.70 per ton, in 1901 for $5.81, in 1916 for 85.92. When the United States entered World War I, flyorspar production and price skyrocketed; the production reached 263,817 tons in 1918, with prices above $20.00 per ton, and prices were even higher in 1919 and 1920. After 1920, the domestic production fell from the 1918 level to about 125,000 tons annually and the average price dropped back to about $18.00 per ton. By 1935 the average price had fallen to $15.00, but in 1940 with a rising demand the average price of all grades had gone back up to $20.31. Today it is 75 per cent higher than in 1940 and presumably the price of acid grade at .least would be even higher if it were not for price ceilings. The first subsequent year that the domestic production exceeded that of 1918 was 1941, and the 1944 production is expected to exceed 400,000 tons, or over one and a half times that of 1918. In five-year periods, average production and imports were as are shown in Table I. Many people concerned with the consumption of fluorspar believe that the postwar period ahead will show a repetition of price trends of the last such period. They believe that imports will be again an important factor in the available supply and the price will fall back gradually toward the prewar level. The following basic factors are pertinent in analyzing this conclusion:

Jan 1, 1948

By J. L. Gillson

In a brief summary of the many occurrences of fluorspar in our western states, it is not possible to go into detail in regard to the geology, mining and milling methods, and reserves about individual deposits. Rather a broad review is made of the occurrences and an analysis of how the western deposits may fit into our future domestic economy is presented. General Statement on Fluorspar Situation Fluorspar has been used for years as a flux in the manufacture of basic open-hearth steel, in foundry cupolas for melting pig iron, in the manufacture of opalescent glass and in the preparation of hydrofluoric acid. In 1880 the annual production of fluorspar in the United States was around 4000 tons per year. By 1900 it was up to 21,656 tons and between 1902 and 1908 it averaged about 40,000 tons annually. Prices were then so low, by present standards, that it seems impossible to believe now that mining could have been performed profitably at the price levels then prevailing. In 1890 fluorspar sold at the mines for $6.70 per ton, in 1901 for $5.81, in 1916 for 85.92. When the United States entered World War I, flyorspar production and price skyrocketed; the production reached 263,817 tons in 1918, with prices above $20.00 per ton, and prices were even higher in 1919 and 1920. After 1920, the domestic production fell from the 1918 level to about 125,000 tons annually and the average price dropped back to about $18.00 per ton. By 1935 the average price had fallen to $15.00, but in 1940 with a rising demand the average price of all grades had gone back up to $20.31. Today it is 75 per cent higher than in 1940 and presumably the price of acid grade at .least would be even higher if it were not for price ceilings. The first subsequent year that the domestic production exceeded that of 1918 was 1941, and the 1944 production is expected to exceed 400,000 tons, or over one and a half times that of 1918. In five-year periods, average production and imports were as are shown in Table I. Many people concerned with the consumption of fluorspar believe that the postwar period ahead will show a repetition of price trends of the last such period. They believe that imports will be again an important factor in the available supply and the price will fall back gradually toward the prewar level. The following basic factors are pertinent in analyzing this conclusion:

Jan 1, 1948

By P. W. Bennett, E. M. Elkin

A new technique for the recovery of palladium and platinum and sludge from go12 electrolyte eliminates many of the drawbacks of the zinc-dust cementation process. In the electrolytic refining of gold by the Wohwill method, the electrolyte contains about 100 gpl gold as AuCl, and 100 gpl free hydrochloric acid. Any palladium and platinum contained in impure gold dissolve and accumulate in the electrolyte. When their concentration approaches that of gold, it becomes desirable to discard the electrolyte and to recover the gold and the palladium and platinum. As practiced at Canadian Copper Refiners until recently, discarded electrolyte was treated with sulfur dioxide to precipitate the gold. 2AuCl3 + 6H2O + 3SO2 —2Au + 6HC1 + 3H2SO4 The gold precipitate was filtered off, washed with water, and melted with scrap anodes. The wash water was added to the gold-free filtrate and treated with zinc dust to cement the palladium and platinum. PdCl2 + Zn — Pd + ZnCl2 PtCl4 + 2Zn—Pt + 2ZnCl2 The precipitate was separated and sold as Pd-Pt concentrate, and the barren filtrate and wash water were discarded. The receipts of palladium and platinum at this company were too small to warrant their refining. The process suffered from a number of disadvantages. Handling the gold-free solution was disagreeable, as it was laden with sulfur dioxide even after blowing with air overnight. Subsequent treatment with zinc dust was also troublesome. A large excess of zinc dust was required. Spray, acid mist, and vigorous evolution of hydrogen accompanied the reaction. The Pd-Pt precipitate was grossly contaminated with occluded unreacted zinc and with yellow granules of sulfur, thought to be formed by the reduction of sulfur dioxide and its derivatives. Repeated boiling of the precipitate with hydrochloric acid and washing with water decreased somewhat, but did not eliminate, the zinc. After drying overnight, the precipitate hardened to a clinkerlike mass that was difficult to grind. Grinding, handling, and packing created heavy dusting. Even weighing in an analytical balance generated dust, indicating that the powder was easily charged with static electricity. The total palladium and platinum content of the powder varied from 56 to 72 pct, with an average of 65 pct. The powder dissolved in aqua regia with difficulty, leaving a small amount of insoluble residue. However, the main drawback of the process was the strongly hygroscopic Pd-Pt product. The powder showed a marked though erratic increase in weight and a corresponding decrease in assay on exposure to air, even during the course of laboratory analysis. Such changes were the source of disagreement between Canadian Copper Refiners and the buyers of concentrate. The cause of the change in weight is believed to be hydrogen, liberated by the action of zinc dust on hydrochloric acid. Some of this hydrogen was adsorbed so strongly by the precipitate that it could not be driven off during the drying. On exposure to air it oxidized catalytically, forming water. Experiments soon showed that the entire process of treatment of discarded electrolyte then in use would have to be changed. A small quantity of apparatus was moved into the gold room, and experimentation was held to a minimum compatible with the development of a new process. PALLADIUM AND PLATINUM PRECIPITATION It was thought that zinc could be replaced with a reagent other than a hydrogen-evolving metal. Semiquantitative experiments showed that platinum and palladium could be precipitated completely from the gold-free solutions with such well-known reagents as sodium formate, formic acid, formaldehyde, and paraformaldehyde. Sodium formate was the preferred reagent, since the other three are volatile and somewhat toxic. The gold-free solutions used were those saturated with sulfur dioxide, partly freed of sulfur dioxide by aerating overnight, completely freed by boiling, as well as strongly and weakly acidic, neutral, and alkaline solutions. The reactions are complex and can be represented by these over-all equations: PdC1, + HCOONa — Pd + NaCl + HC1 + Co2 PtCl4 + 2HCOONa — Pt + 2NaCl + 2HC1 + 2CCX, Strongly acidic solutions required as much as five times the theoretical amount of sodium formate as did solutions neutralized to pH 4.5. Increasing the pH decreased the consumption of

Jan 1, 1965

By Attilio Ligasacchi

The production of magnetite in copper smelting has been the subject of wide investigation in the past.1-9 It is accepted that most of the magnetite is produced in the converters and returned to the re-verberatory furnace with the converter slag. Part of this magnetite is circulated back to the converters in the matte. As a result, if proper fluxing practice is not followed, and temperatures in the converting process are not strictly controlled, an undesirable circulating load of magnetite may occur. This might ultimately lead to saturation in the reverberatory bath with consequent settling of the magnetite at the furnace bottom. This paper contains the cumulative results of a series of investigations on the formation and control of magnetite presently being conducted by the metallurgical department of the Chino Mines Division of Kennecott Copper Corp. THE METHOD USED FOR THE DETERMINATION OF MAGNETITE AND OTHER MAGNETIC MATERIALS Conventional wet chemical magnetite analyses are neither reproducible nor theoretically sound. These analyses are sensitive to time, temperature, and heating rate-all factors which tend to decrease their accuracy and reproducibility. To overcome the unreliability of the above method, H. W. Mossman1 developed an inductance bridge magnetite "meter" which completely replaces the wet chemical analysis with the advantage of quicker and reproducible results. Fig. 1 is a schematic diagram of the meter in use in the Chino metallurgical laboratory. When a pulverized specimen with magnetic properties is placed in coil A of the inductance bridge circuit, the permeability of the core is increased, causing a voltage drop across the coil. The unbalance is amplified to operate the balancing motor which adjusts the movable iron core in coils C and D to bring the bridge back to balance. An indicator arm connected to the movable iron core indicates directly the percent, or weight, of the magnetic material in the specimen. Coil B acts as a reference coil and a zero adjustment. The meter is standardized with a set of various amounts of -100 mesh pure magnetite crystals (99.8 pet Fe3O4) thoroughly mixed in a test tube with -100 mesh quartz sand. These mixtures are combined with inert transoptic powder and heated to 140° to 160°C (melting point of the powder) and subsequently cooled to room temperature, solidifying the whole mixture, in this manner eliminating any possible preferential orientation of the magnetite grains. In order to determine the effect of the particle size" of the magnetite and other magnetic materials on the meter determinations, five size fractions were obtained by wet screening of the 99.8 pet pure magnetite and several reverberatory matte and slag samples. The following size fractions were investigated: -80 + 100; -100 + 150; -150 + 180; -200 + 250; -250 + 300 and -100 mesh. The deviation of the meter readings for the various sizes of each sample never exceeded ±0.3 pet. This deviation is considered well within the accuracy needed for common smelter operation procedures. Further investigation proved also that grinding slag and matte samples in our laboratory disc grinder normally produced a powder constituted of 95 pet -200 + 300 mesh. For practical purposes a - 100 mesh size was used as standard for our studies. It was also found that for rapid determinations tamping of the sample in the test tube was sufficient and gave comparable results with those samples solidified with transoptic powder. For the purpose of this study we define the magnetite content in copper smelter slags and matte as the equilibrium amount of crystalline Fe3O4 existent in the specimen at a given reference temperature, excluding any secondary magnetite precipitated in the specimen during nonequilibrium conditions. MAGNETITE DETERMINATION IN SLAGS Since it is impossible to ascertain the status of equilibrium of a conventional slag sample, due to

Jan 1, 1965

By J. E. Hanlon, S. R. Butler, R. J. Wasilewski

It is known that stoichiometric NiTi transforms to a structure of lower symmetry near room temperature. The present investigation deals primarily with the changes in the electrical and magnetic properties which occur at the transformation. The effects of composition and various heat treatments on electrical properties are also discussed. In recent investigations by Purdy and Parr,' Wang et a1.,2 Dautovich and Purdy,~ Dautovich et a1.,4 and the authors, ~,' NiTi has been observed to undergo a martensitic phase transformation from the ordered CsCl (B2) structure to one of lower symmetry. The crystal structure of the low-temperature modification is still not completely worked out. This transformation occurs near room temperature for the stoichiometric composition. The thermodynamic, elastic, crystallographic, and microstructural characteristics of this transformation for some compositions have been previously reported.'-= In the present investigation, the changes in the electrical transport properties and magnetic susceptibility of NiTi when it undergoes the martensitic phase transformation have been measured. The approximate composition range covered is 48 to 52 pct Ti.* *All compositions are given in atomic percent._____________________ EXPERIMENTAL PROCEDURES AND RESULTS 1) Sample Preparation. The NiTi alloys were prepared from iodide titanium (>99.92 pct Ti) and vacuum-premelted carbonyl nickel (>99.95 pct Ni) by repeated arc melting under gettered argon and casting in a water-cooled copper mold. After the cast buttons were hot-forged at 800" to 950°C, the oxidized surface was removed by grinding and etching. Samples were cut from the worked polycrystalline buttons. After fabrication, the samples were annealed 1 hr at 900°C and water-quenched. The composition of the alloys was assumed to be the same as in the initial feed. This assumption was supported by material balance checks on 100-g arc-melted buttons which demonstrated quantitative recovery to within k0.02 pct. The oxygen content of the alloys was '150 ppma (weight) and the carbon content was 30 to 50 ppm. 2) Electrical Resistivity and Hall Effect. The electrical resistivity was measured over the composition range 48 to 52 pct Ti for temperatures varying from 4.2"K to 900°C. The high-temperature and many of the low-temperature experiments were carried out using conventional 4-probe dc techniques.7 The Hall coefficient and ac resistivity were measured at 20 cps using the apparatus shown schematically in Fig. 1. The equipment is basically an ac potentiometer: The resistive or Hall voltage of the sample is balanced by a comparison voltage provided by a dividing network across a standard series resistance. The off-balance signal is amplified by a Tektronix 122 amplifier and detected by a Princeton Applied Research (PAR) phase-sensitive detector. The precision of the resistivity and Hall coefficient measurements is i0.1 and i3 pct, respectively. The uncertainty in the data is increased by i2 pct from the precision limits because of uncertainty in dimensions and possible end effects. The temperature dependence of the electrical resistivity and Hall effect for a stoichiometric (50.0 pct Ti) NiTi sample are illustrated in Fig. 2. Above 75"C, this sample has essentially the B2 structure, while below about 0°C it is completely in the low-tempera-

Jan 1, 1968

On June 15, Dr. E. E. Southard gave the following report to the Advisory Committee on Industrial Mental Hygiene of the. Engineering Foundation. Psychiatric social work, a new specialty in social work, was the topic of several section meetings (500-600 in attendance) at the National Conference for Social Work, held at Atlantic City, June 1-6. The new specialty was stated, by Dr. R. C. Cabot, to be the greatest single advance in medical social work since its establishment. Reports from three schools (Smith, New York, Philadelphia) indicate a good production rate of personnel, probably at least one hundred a; year, chiefly to be absorbed by hospitals, unless industrial psychiatry steps into the field. No detailed public announcements concerning industrial psychiatry were made; but private discussions elicited approval only for the program. Indications are that many non-medical social workers will decide to get additional training to go into medical social work and then to specialize in psychiatric social work and that many medical social workers will desire to specialize in psychiatric social work, with a view partly to entering reconstruction work, partly to entering industrial psychiatry. Suggestion has been made that psychiatric social workers may go into some kinds of, employment management and may look for special courses to that end. Evidence. has been brought that certain "compensated" psychopathic employees have been good (even very good) workmen: emphasis should be laid on this point to show the constructive aspect of industrial psychiatry:

Jan 8, 1919

By C. A. Rowland, R. C. Nealey

Until recently there were very few dry grinding rod mills used. The early installations performed as anticipated. In the last seven years a number of dry grinding rod mills have been made, most of which have been used for grinding iron ore. The recent installations have not operated as anticipated in accordance with practices learned from wet rod milling. The paper does not establish any definite rules for the application of dry grinding rod mills but does attempt to establish some guidelines that can be followed. The paper reviews a number of flowsheets describing the problems encountered and some of the solutions found. Completely dried compared to partially dried feed is reviewed. Problems encountered in scaling up from pilot plant to production-size mills are described. Rod sizing, effect of rod length, rod bending, rod breakage, and rod tangling are discussed. While wet grinding rod mills have been used for 50 years to grind ores and make specifications sands, dry grinding rod mills have only recently come into significant use in the mineral industries. Among the materials now being processed in dry grinding rod mills are (1) lead sinter, lead slag, and limestone; (2) iron ore; (3) silica; (4) cement clinker; and (5) coke. Because the application of wet grinding rod mills was well understood at the time that dry grinding rod mills were introduced, the criteria and guidelines for selecting the former were carried over to the selection of the latter - in some cases with the anticipated results, but in many cases not. It is the purpose of this paper to describe, through a series of flowsheets, some of the recent problems encountered with dry grinding rod mills, to point out the solutions to these problems, and to set forth some guidelines for the selection and application of these mills. When dry grinding rod mills came into use, the following facts about their wet grinding counterparts were already well established: (1) The performance of large-sized mills can be accurately predicted from the performance of rod mills 5 or 6 ft in diameter. Wet grinding mills as large as 14 ft in diameter (inside shell) by 20 ft long with rods 19.5 ft long can operate satisfactorily.

Jan 1, 1970

By C. R. Holmgren

A large consolidated sand core was used in this investigation. Four groups of experiments including gas drives, water drives, and water drives combined with gas injection, were made. The results of the gas-drive experiments indicate that greater ultimate recovery may be obtained by higher pressure gradients. However, other considerations challenge the absolute veracity of this conclusion. Results indicate that with connate water present and a low pressure gradient during a gas drive, less gas is required to produce a given amount of oil. The saturation gradient determined during a gas drive is presented. Results of the water-drive experiments indicate that the recovery of oil is not a function of either gradient Or input rate. The presence of initial gas saturation resulted in a slightly lower final oil saturation. Injection of sufficient gas to maintain the initial gas saturation during a water drive resulted in lower final oil saturation. Introduction With the demand for petroleum products reaching an all-time record high and the location and exploitation of new reserves becoming more difficult and expensive, the oil industry must look to more thorough and efficient methods of producing the petroleum already located. Much of the previous work along these lines is exploratory and suggests a more complete investigation. For example, a great amount of laboratory work has been done on flooding sands initially saturated roo pct with oil. Since sands saturated roo pct with oil rarely, if ever, exist in nature, and because the water content is a vital factor in the ultimate recovery from the sand, laboratory investigations on cores without water must be supplemented by comparable work on cores with water present.1'2 Other investigators have done a considerable amount of laboratory work using gas and water drives.3 Krutter and Day4 have reported results of investigations using cores saturated IOO pct with oil and have concluded that " the ultimate recovery obtained, using air drive, from a completely oil-saturated core depends on the pressure gradients used. The higher the pressure gradient, the greater will be the recovery until a certain Pressure range is reached above which the reward of increased recovery is much smaller for similar increments in pressure." Yuster and Day6 conclude that less gas and less time are required to produce a given quantity of oil at high pressure gradients than at low pressure gradients; they also conclude that an increase in the viscosity of the oil requires a proportionate increase in the gas needed for recovery. Their results indicate that the gas required to produce a given quantity of oil is greatly reduced by the presence of connate water. R. C. Earlougher2 concluded, based on results of water-flooding a large number of

Jan 1, 1949

By G. Y. Chin, A. R. Von Neida

The relationship between crystalline texture and magnetic anisotropy in a Co-10 pct Fe alloy has been investigated by means of X-ray pole .figures, hysteresis-loop tracings, and magnetic-torque measurements. The cold-rolled texture of a sample after 95 pct thickness reduction is of the "brass-type", that is, (110)[112/. At 500°C the cold-rolled material recrystallizes fo a (311)[112] orientation, plus a minor (210) [001] component. This texture persists up to 1400°C, at which point it is replaced by a secondary recrystallization texture, (111)[112]. Magnetic-torque measurements show that the easy axis of magnetization lies along the transverse direction in the cold-rolled material and changes to the rolling direction after primary recrystallization. This change is also reflected in the hysteresis characteristic; whereas the B-H loop along the rolling direction is nonsqunre with a low value of residual induction (750 vs 19,000 gauss at saturation) for the cold-rolled material, the annealed sample exhibits a very square loop with a large value of residual induction (16,000 gauss). Theoretical torque curves calculated from the texture data on the basis of magnetocystalline anisotropy energy compare favorably with the observed torque curves. Values of the crystal anisotropy constant, K1, obbained from these comparisons are in good quantitative agreement with single-crystal data. Furthermore, both the torque amplitude of the poly crystalline material and the magnitude of K1 obtained with a single crystal show an identical temperature dependence over the range from —195° 10 +220°C. It is thus conclzided that the magnetocrystalline anisotrpopy erlergy is chiefly I-esponsible. for the nzagnetic allisotropy of textured Co-10 pct Fe alloy. THE results of a study on the relationship between crystalline texture and magnetic anisotropy of an alloy containing 90 pct Co and 10 pct Fe by weight are reported in this paper. The alloy has an fcc structure: a saturation magnetization of about 19,000 gauss,2 and a nearly zero value of magnetoatriction.3 puzey4 has reported a large crystalline anisotropy constant of -7.4 X 105 ergs per cu cm, indicating the (111) directions as the easy axes of magnetization. Hence, texture variations resulting from cold work and annealing are expected to introduce a magnetic anisotropy in the polycrystalline material through the alignment of these (111) easy directions. In the present study, magnetic anisotropy of both rolled and annealed strips was determined by magnetic-torque measurements and hysteresis-loop tracings The results were correlated quantitatively with calculations based on texture data obtained by the X-ray pole-figure technique. It is established that the magnetocrystalline anisotropy energy is indeed chiefly responsible for the magnetic anisotropy of textured Co-10 pct Fe alloy. EXPERIMENTAL Material Preparation. The alloy nominally contained 90 pct Co, 9.5 pct Fe, and 0.5 pct Mn by weight, the chief impurities being nickel, carbon, and silicon (less than 0.5 pct total). The cast ingot was rolled to strips 10 mils thick and slit to 1 in. in width. After strand annealing at 900°C in a hydrogen atmosphere, these strips were further cold-rolled to various thicknesses down to 0.25 mil (97.5 pct reduction), with samples collected for study of the development of texture and magnetic anisotropy as a function of thickness reduction. subsequently, samples cold-rolled from 100 mils to 0.25 mil were obtained, thus extending the thickness reduction range to 99.75 pct. In addition, 0.5-mil samples cold-rolled from 10 mils were selected

Jan 1, 1965

By H. H

This paper describes the milling practice and operating results of the recently installed heavy-media separation plant of Barton Mines Corporation, the world's largest producer of garnet. This pioneer application of heavy-media separation (sink-float) processes for the beneficiation of a nonmetallic ore is unique in many respects, the details of which will be apparent from the following presentation. Characteristics of the Ore The property is on Gore Mountain, 5 miles west of the village of North Creek, New York, and 11 miles by road. The altitude of the mine is approximately 2800 feet. The valuable mineral, garnet, occurs in a surface deposit, the ore body being about ¾, mile long and varying in width from 50 to 300 ft. The garnet, principally almandite, Fe3Al2 (SiO4)3, occurs in a metamorphic rock of uncertain origin. The gangue mineral is principally hornblende, which constitutes from 40 to 80 per cent of the rock. The remainder of the gangue is divided between plagioclase feldspars and hypersthene, with smaller amounts of biotite, apatite and pyrite. The garnet content of the ore body varies from 5 to 20 per cent of the mass and averages about 10 to I2 per cent. The garnet occurs as crystals, mostly imperfectly developed, known locally as "pockets." These crys- tals vary in size from a fraction of an inch to a foot or more in diameter. Occasionally crystals have been found up to 30 and 36 in. in diameter but they average 4 to 6 inches. Nearly every crystal of garnet is surrounded by a rim of coarsely crystalline hornblende. The quarry faces are striking in appearance, showing crimson-red garnet crystals with their coal black rims scattered over a grayish black background. The specific gravity of the garnet is 3.8 to 4.1 while that of the hornblende ordinarily is 3-07 to 3.24. Some specimens of very dense hornblende have been found with a specific gravity as high as 3.40. The garnet crystals are laminated in structure, and readily cleave into flat, tabular pieces with sharp, chisel-like edges. Even when crushed to a very fine size, the garnet retains this flat, sharp, slivery grain shape. This characteristic, together with its hardness and toughness, gives it its value as an abrasive- Former Operations Mining operations have been carried on for 60 years. Until 1923, only the oxidized (soft rock) section of the ore body was mined. The oxidation, which affected the garnet only slightly, made possible the removal of the garnet from the rock by hand after it was blasted down, using small hand picks. In 1923 a mill was erected on the property, and was rebuilt and redesigned in 1928. Before the adoption of heavy-media separation, concentration was effected by

Jan 1, 1943

By H. H

This paper describes the milling practice and operating results of the recently installed heavy-media separation plant of Barton Mines Corporation, the world's largest producer of garnet. This pioneer application of heavy-media separation (sink-float) processes for the beneficiation of a nonmetallic ore is unique in many respects, the details of which will be apparent from the following presentation. Characteristics of the Ore The property is on Gore Mountain, 5 miles west of the village of North Creek, New York, and 11 miles by road. The altitude of the mine is approximately 2800 feet. The valuable mineral, garnet, occurs in a surface deposit, the ore body being about ¾, mile long and varying in width from 50 to 300 ft. The garnet, principally almandite, Fe3Al2 (SiO4)3, occurs in a metamorphic rock of uncertain origin. The gangue mineral is principally hornblende, which constitutes from 40 to 80 per cent of the rock. The remainder of the gangue is divided between plagioclase feldspars and hypersthene, with smaller amounts of biotite, apatite and pyrite. The garnet content of the ore body varies from 5 to 20 per cent of the mass and averages about 10 to I2 per cent. The garnet occurs as crystals, mostly imperfectly developed, known locally as "pockets." These crys- tals vary in size from a fraction of an inch to a foot or more in diameter. Occasionally crystals have been found up to 30 and 36 in. in diameter but they average 4 to 6 inches. Nearly every crystal of garnet is surrounded by a rim of coarsely crystalline hornblende. The quarry faces are striking in appearance, showing crimson-red garnet crystals with their coal black rims scattered over a grayish black background. The specific gravity of the garnet is 3.8 to 4.1 while that of the hornblende ordinarily is 3-07 to 3.24. Some specimens of very dense hornblende have been found with a specific gravity as high as 3.40. The garnet crystals are laminated in structure, and readily cleave into flat, tabular pieces with sharp, chisel-like edges. Even when crushed to a very fine size, the garnet retains this flat, sharp, slivery grain shape. This characteristic, together with its hardness and toughness, gives it its value as an abrasive- Former Operations Mining operations have been carried on for 60 years. Until 1923, only the oxidized (soft rock) section of the ore body was mined. The oxidation, which affected the garnet only slightly, made possible the removal of the garnet from the rock by hand after it was blasted down, using small hand picks. In 1923 a mill was erected on the property, and was rebuilt and redesigned in 1928. Before the adoption of heavy-media separation, concentration was effected by

Jan 1, 1943

By F. C. von der Weid

The mining district of Sao Fidelis, in the northern part of the State of Rio de Janeiro, is situated on the Paraiba River, where it crosses the Serra do Mar (Mountain of the Sea), in the center of a mountainous country having numerous graphitic outcrops. For a long time, these outcrops were irregularly explored by local people, who had little means and no technical advice. A few years ago, the Cia. Brasileira de Minerasgo de Grafite took over the control of some of the more important outcrops and entrusted to me the task of studying the ores and establishing a flowsheet for a pilot plant to be situated in São Fidelis—the ore to be concentrated being mostly of a very high grade. Mineralization and Ores The rocks of the district are mostly gneiss and crystalline schists, with occasional interstratification of dolomitic limestone belonging to the Archean Brazilian Fundamental Complex. The general direction of the beds is roughly between 60" and 80°E., with variable dip, generally nearly vertical. Mostly, direction and dip of the graphitic outcrops is concordant with the embanking gneisses or schists. These outcrops seem to be aligned in two or three parallel zones (Pádua, São Fidelis, Mac- abu), the Sao Fidelis zone being the best known and the most important to date. Two outcrops of the São Fidelis zone, Saudade and Sao Benedito, are now being exploited. Mining has been carried out, so far, on a small scale. The mineralization consists of a vein of nearly massive graphite in a very compact, partially granitized gneiss, with vertical dip. The vein is most irregular, with nests and lenses irregularly distributed along the fracture plane. The ore consists chiefly (particularly in Sao Benedito) of massive graphite appearing and disappearing along the main direction of the vein, forming a succession of irregular lenses in a sort of rosary design. The average width is more or less one foot, with variations from zero to 6 feet. The lode minerals are quartz, a little feldspar and garnet (the latter having its origin perhaps in the embedding gneisses), with rare ferruginous clays on the outcrops. In Saudade a little pyrite, pyrrhotite and chalcopyrite is found occasionally, as well as a carbonaceous mineral, very light, black and shiny, with conchoidal fracture, looking like a sort of jet. The composition of this mineral has not yet been established, but it seems to be very complex. The mineralization seems to be hydro-thermal at a rather low temperature, perhaps mesothermal. The paragenesis, according to samples from Saudade, seems to be as follows, the earlier minerals being listed first: pyrrhotite, pyrite, chalcopyrite, with quartz concomitant to the three and graphite formed later. Age of the mineralization is still uncertain, owing to the very in-

Jan 1, 1948

By F. C. von der Weid

The mining district of Sao Fidelis, in the northern part of the State of Rio de Janeiro, is situated on the Paraiba River, where it crosses the Serra do Mar (Mountain of the Sea), in the center of a mountainous country having numerous graphitic outcrops. For a long time, these outcrops were irregularly explored by local people, who had little means and no technical advice. A few years ago, the Cia. Brasileira de Minerasgo de Grafite took over the control of some of the more important outcrops and entrusted to me the task of studying the ores and establishing a flowsheet for a pilot plant to be situated in São Fidelis—the ore to be concentrated being mostly of a very high grade. Mineralization and Ores The rocks of the district are mostly gneiss and crystalline schists, with occasional interstratification of dolomitic limestone belonging to the Archean Brazilian Fundamental Complex. The general direction of the beds is roughly between 60" and 80°E., with variable dip, generally nearly vertical. Mostly, direction and dip of the graphitic outcrops is concordant with the embanking gneisses or schists. These outcrops seem to be aligned in two or three parallel zones (Pádua, São Fidelis, Mac- abu), the Sao Fidelis zone being the best known and the most important to date. Two outcrops of the São Fidelis zone, Saudade and Sao Benedito, are now being exploited. Mining has been carried out, so far, on a small scale. The mineralization consists of a vein of nearly massive graphite in a very compact, partially granitized gneiss, with vertical dip. The vein is most irregular, with nests and lenses irregularly distributed along the fracture plane. The ore consists chiefly (particularly in Sao Benedito) of massive graphite appearing and disappearing along the main direction of the vein, forming a succession of irregular lenses in a sort of rosary design. The average width is more or less one foot, with variations from zero to 6 feet. The lode minerals are quartz, a little feldspar and garnet (the latter having its origin perhaps in the embedding gneisses), with rare ferruginous clays on the outcrops. In Saudade a little pyrite, pyrrhotite and chalcopyrite is found occasionally, as well as a carbonaceous mineral, very light, black and shiny, with conchoidal fracture, looking like a sort of jet. The composition of this mineral has not yet been established, but it seems to be very complex. The mineralization seems to be hydro-thermal at a rather low temperature, perhaps mesothermal. The paragenesis, according to samples from Saudade, seems to be as follows, the earlier minerals being listed first: pyrrhotite, pyrite, chalcopyrite, with quartz concomitant to the three and graphite formed later. Age of the mineralization is still uncertain, owing to the very in-

Jan 1, 1948

By A. T. Sweet, C. E. Plummer, H. W. St. Clair, S. F. Ravitz

The ammonium sulphate process for recovering alumina from clays was proposed by Rinman, Buchner, and others many years ago, and more recently various modifications have been investigated both here arid abroad.l-Is The process, as it is being investigated by the Bureau of Mines, involves the following essential steps: (I) sulphating the alumina in the clay by baking with ammonium sulphate at approximately 4m°C., (2) leaching with water to extract aluminum sulphate and ammonium sulphate, (3) crystallization of ammonium alum, (4) conversion of the alum to aluminum hydroxide by the ammonia gases given off during the baking, and (5) calcining the aluminum hydroxide to alumina. The over-all reaction between kaolinite (the most important alumina mineral in most clays) and ammonium sulphate during the baking step may be expressed as follows: Al2O8.2SiO2.2H2O + 4(NH4)2SO4 = (NH4)2SO4.A12(SO4)3 + 6NH3(g) + SH2O(g) + 2SiO2 [i] There is definite evidence. as is shown later. that the aluminum occurs in the baked product as anhydrous ammonium alum rather than as aluminum sulphate, provided the baking temperature is not much above 4o0°C. It is probable that the reaction takes place in the following steps: (NH4)2SO4 = NH3 -(- NH4HSO4 [2] Al2O3.2SiO2.2H2O + ;}NH4HSO4 = A12(SO4)3 + 3NH3 + 5H2O + 2SiO2 [3] A12(SO4)3 + (NH4)2SO4 = (NH4)2SO4.A12(SO4)3 [4] In onc modification of the process, the ammonium sulphate is first decomposed according to Eq. z, and the clay is treated with the resultant ammonium acid sulphate. Reaction I is highly endothermic. On the basis of the best available data, the following values have been estimated for the changes in heat content and free energy in calories per mol: -AH = 331,040 - 34T -AF = 331~040 - 78.3T log T - 671.8T The heat required' at 25OC is 320,900 cal. per mol, or 5650 B.t.u. per pound of alumina. The theoretical equilibcium pressure of NHa, based on the free-energy equation given above, increases very rapidly as the temperature rises above 3o0°C., and reaches one atmosphere at 43s°C. Ammonium alum solutions are almost ideal for crystallization. A saturated solution of ammonium alum contains 126. grams of AlzOa per 1000 grams of water at 90°C. and only 16.5 at 20 C. Hence most of the alum crystallizes out when the

Jan 1, 1944

By H. H. Lowry

SCIENCE attracts the attention and interest of an individual or an industry in general only in proportion to the apparent direct application to its immediate welfare or benefit. Engineering accomplishments, as applied science, therefore receive more recognition than the scientific discoveries on which they are based. Even self-evident truth needs endless repetition before recognition: engineering as applied science cannot advance beyond the science on which it is founded. Scientific research, and discovery, is sometimes forced by the limitations recognized by those active in its development for the use of society and industry and sometimes results from an apparently superficially illogical curiosity about natural phenomena. Research has received inadequate support except in industries born in the laboratory, the chemical and electrical industries, from which many lessons might well be drawn by their elder associates. The rapid growth of these industries has been largely due to close cooperation of scientists and engineers. Most rapid advance is made when the research worker and the engineer keep in close touch with each other's ideas and problems. Many of the developments of modern industry are based on dis-coveries buried for years in scientific journals because the "time was not ripe" at their discovery. The situation with regard to utilization of coal is quite the opposite. The fundamental facts in regard to the scientific aspects of the problem of developing the energy from coal are scarcely known. The engineer, without much scientific cooperation, has accom-plished much in increasing the number of kilowatt-hours per pound of coal from 0.31 in 1919 to 0.69 in 1934, in increasing the 1000 gross ton-miles of freight hauled from 6.1 to 8.3 per ton of coal in the same period, in fuel economy in production of coke, and in developing the production of motor fuel in Germany and England by hydrogenation of coal. This lack of scientific research on the utilization of our most important natural resource, coal, has been due undoubtedly in large part to the reluctance of investigators to study the interrelationships involved in, the complicated processes in which coal is used. These complexities are made

Jan 1, 1936

By F. C. von der Weid

The mining district of Sao Fidelis, in the northern part of the State of Rio de Janeiro, is situated on the Paraiba River, where it crosses the Serra do Mar (Mountain of the Sea), in the center of a mountainous country having numerous graphitic outcrops. For a long time, these outcrops were irregularly explored by local people, who had little means and no technical advice. A few years ago, the Cia. Brasileira de Mineracao de Grafite took over the control of some of the more important outcrops and entrusted to me the task of studying the ores and establishing a flowsheet for a pilot plant to be situated in Sao Fidelis—the ore to be concentrated being mostly of a very high grade. Mineralization and Ores The rocks of the district are mostly gneiss and crystalline schists, with occasional interstratification of dolomitic limestone belonging to the Archean Brazilian Fundamental Complex. The general direction of the beds is roughly between 60' and 80°E., with variable dip, generally nearly vertical. Mostly, direction and dip of the graphitic outcrops is concordant with the embanking gneisses or schists. These outcrops seem to be aligned in two or three parallel zones (Padua, Sao Fidelis, Mac- abu), the Sao Fidelis zone being the best known and the most important to date. Two outcrops of the Silo Fidelis zone, Saudade and Sao Benedito, are now being exploited. Mining has beer] carried out, so far, on a small scale. The mineralization consists of a vein of nearly massive graphite in a very compact, partially granitized gneiss, with vertical dip. The vein is most irregular, nith nests and lenses irregularly distributed along the fracture plane. The ore consists chiefly (particularly in Sao Benedito) of massive graphite appearing and disappearing along the main direction of the vein, forming a succession of irregular lenses in a sort of rosary design. The average width is more or less one foot, with variations from zero to 6 feet. The lode minerals are quartz, a little feldspar and garnet (the latter having its origin perhaps in the embedding gneisses), with rare ferruginous clays on the outcrops. In Saudade a little pyrite, pyrrhotite and chalcopyrite is found occasionally, as well as a carbonaceous mineral, very light, black and shiny, with conchoidal fracture, looking like a sort of jet. The composition of this mineral has not yet been established, but it seems to be very complex. The mineralization seems to be hydro-thermal at a rather low temperature, perhaps mesothermal. The paragenesis, according to samples from Saudade, seems to be as follows, the earlier minerals being listed first: pyrrhotite, pyrite, chalcopyrite, with quartz concomitant to the three and graphite formed later. Age of the mineralization is still uncertain, owing to the very in-

Jan 1, 1947

By F. C. von der Weid

The mining district of Sao Fidelis, in the northern part of the State of Rio de Janeiro, is situated on the Paraiba River, where it crosses the Serra do Mar (Mountain of the Sea), in the center of a mountainous country having numerous graphitic outcrops. For a long time, these outcrops were irregularly explored by local people, who had little means and no technical advice. A few years ago, the Cia. Brasileira de Mineracao de Grafite took over the control of some of the more important outcrops and entrusted to me the task of studying the ores and establishing a flowsheet for a pilot plant to be situated in Sao Fidelis—the ore to be concentrated being mostly of a very high grade. Mineralization and Ores The rocks of the district are mostly gneiss and crystalline schists, with occasional interstratification of dolomitic limestone belonging to the Archean Brazilian Fundamental Complex. The general direction of the beds is roughly between 60' and 80°E., with variable dip, generally nearly vertical. Mostly, direction and dip of the graphitic outcrops is concordant with the embanking gneisses or schists. These outcrops seem to be aligned in two or three parallel zones (Padua, Sao Fidelis, Mac- abu), the Sao Fidelis zone being the best known and the most important to date. Two outcrops of the Silo Fidelis zone, Saudade and Sao Benedito, are now being exploited. Mining has beer] carried out, so far, on a small scale. The mineralization consists of a vein of nearly massive graphite in a very compact, partially granitized gneiss, with vertical dip. The vein is most irregular, nith nests and lenses irregularly distributed along the fracture plane. The ore consists chiefly (particularly in Sao Benedito) of massive graphite appearing and disappearing along the main direction of the vein, forming a succession of irregular lenses in a sort of rosary design. The average width is more or less one foot, with variations from zero to 6 feet. The lode minerals are quartz, a little feldspar and garnet (the latter having its origin perhaps in the embedding gneisses), with rare ferruginous clays on the outcrops. In Saudade a little pyrite, pyrrhotite and chalcopyrite is found occasionally, as well as a carbonaceous mineral, very light, black and shiny, with conchoidal fracture, looking like a sort of jet. The composition of this mineral has not yet been established, but it seems to be very complex. The mineralization seems to be hydro-thermal at a rather low temperature, perhaps mesothermal. The paragenesis, according to samples from Saudade, seems to be as follows, the earlier minerals being listed first: pyrrhotite, pyrite, chalcopyrite, with quartz concomitant to the three and graphite formed later. Age of the mineralization is still uncertain, owing to the very in-

Jan 1, 1947