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Papers presented at two joint meetings of the Mineral Industry Education Division, Geophysics Committee, the American Geophysical Union and the Committee on College Curricula of the American Association of Petroleum Geologists, at the A I M E meeting in New York on Feb 17 and 18, 1941 At the first meeting, the Musing Geology Committee, the Society of Economic Geologists and the Committee on Applications of Geology of the American Association of Petroleum Geologists also took part [ ]

Jan 1, 1941

By E. A. Peretti

BEFORE embarking on a detailed investigation of the system indium-antimony-bismuth, experiments were carried out to determine the existence of quasi-binary sections. The bordering binary systems are known1 and show the occurrence of three intermetallic binary species corresponding closely to InSb, InBi and In2Bi. Attention was focused on the low-indium end of the ternary diagram where the possible quasi-binary sections, excluding ternary compound formation, would involve the pairs: InSb-Bi, InBi-Sb and InSb-InBi. The latter pair revealed itself as a true quasi-binary, so it was investigated fully by thermal analysis, x-ray and microscopic examinations. All compositions were made by proper combination of three elements involved. The indium had a guaranteed purity of 99.97 pet; the antimony contained 99.99 pet Sb, and the bismuth had a purity of 99.99 pet. Alloys were made by melting together antimony, bismuth and indium in evacuated, sealed pyrex-glass tubes at a temperature of 600°C with vigorous agitation. The melts, weighing about 80 grams, were allowed to freeze in the sealed tubes and finally were transferred to glass tubes for the taking of cooling and heating curves. A protective atmosphere of helium was used, and the melts were stirred mechanically while cooling at rates of from 0.5 to 3°C per min. Temperatures were measured with a calibrated chromel-alumel thermocouple and were recorded graphically on a Minneapolis-Honeywell extended-range recorder working in conjunction with a Leeds and Northrup type K-2 potentiometer. Conventional polishing techniques gave satisfactory results in preparing specimens for microscopic examination. Etching was carried out with Vilella's reagent or aqueous mixtures of chromic oxide with either nitric acid or sulfuric acid plus dichromate. Conclusions deduced from the thermal analysis and microscopic examinations were confirmed by X-ray photograms using Co-K alpha radiation. The phase diagram is shown in Fig. 1. The liqui-dus consists of an uninterrupted line running from the freezing point of InSb (525°C) to a eutectic point experimentally indistinguishable from the freezing point of InBi (110°C). X-ray patterns of the InSb show a slight expansion of the lattice with the addition of InBi, and microscopic work indicated the solid solubility of InBi in InSb to be less than 0.5 pet at room temperature. Acknowledgment We wish to thank the Indium Corporation of America for the loan of the indium used in this investigation. References 1 E. A. Peretti and S. C. Carapella, Jr.: Trans. ASM., 1949, vol. 41, P. 947. T. S. Liu and E. A. Peretti: Trans. ASM, 1952, vol. 44, p. 539. Metals Reference Book, vol. 1, 2nd ed., AIME, p. 345.

Jan 1, 1959

By Didyk M. Marko, Vásquez Ñ. Gregorio

INTRODUCTION The Salvador Division of Codelco Chile is an important copper molybdenum production center, located 1,200 km north of Santiago and 124 km east of the Pacific Ocean, in the foothills of the Andes Range. El Salvador Mine produces by block caving 8,600,000 t of ore with a content of 1 .l0% Cu and 0.02% Mo, from a copper porphyry ore body. The geology of the ore body shows a great heterogeneity as to rocks, grade distribution, ore height and undercutting levels. This has resulted in a variety of designs for the productive units (blocks) and in the mineral handling systems. Up to now, 4 systems have been employed: grizzly and intermediate haulage; slusher drifts with intermediate ore pass and haulage; slusher drifts with scraper loading directly into trains; and, LHD. On the other hand there is a marked dispersion in the final results from each block mined out. We report here the conclusions reached after analyzing the data gathered on the mining of 163 depleted blocks, representing a total production of 102,000,000 t, with an undercut area of 340,000 sq m. The final results of each block, characterized in terms of tonnage and fine cop- per recovery, were studied as a function of a number of design, geological, and operational parameters. BASIC CONCEPTS Indices of Efficiency In order to evaluate the final result of each block, or group of blocks, the following indices of efficiency were defined : Tonnage Recovery (TR) . It is the ratio between the ore tonnage extracted and the in situ tonnage estimate for each block or group of blocks. Grade efficiency (E) . It is the ratio between the Cu mill feed grade and the in situ Cu grade estimate for each block or group of blocks. Grade efficiency is numerically equal to the ratio between fine copper recovery (FR) and tonnage recovery (TR) . Figure 1 shows the final results for 163 mined out blocks, in terms of tonnage and fine copper recovery. Different grade efficiencies are also indicated. The marked dispersion in the final results of mined out blocks must be stressed since it greatly incides in mine planning. It must be kept in mind that an eventual concentration of low overall efficiency in the blocks during a given period (e.g., one year) will make it necessary to bring into production new blocks over the standard development schedule. It is worth noting that the final results in the 163 mined out blocks corresponding to a production of 102,000,000 t of ore with a mill feed grade of 1.445% were obtained from an ore estimate of 93,000,000 t and a grade of 1.671% Cu, which represents 110% tonnage recovery and 86% grade efficiency. Parameters Studied The following parameters are considered in this study: Geological Parameters. These parameters depend exclusively on the geological and geotechnical features of the ore body. IN SITU GRADE: The initial estimate of the block grade. Four copper grade categories were considered : (a) lower than 1.3% ; (b) 1.3 to 1.6%; (c) 1.6 to 2.0%; (d) higher than 2.0%. The overall initial in situ estimate of the ore body was 1.6% Cu. IN SITU TONNAGE: Initial tonnage estimate of the block. Four categories are considered: (a) below 500,000 t; (b) 500,000 to 1,000,000 t (c) 1,000,000 to 1,500,000 t; (d) over 1,500,000 t. CONTACTS SLOPE: Since up to 4 rock types can be found in a block, resulting in a corresponding number of contacts, a horizontal con- tact was defined as that which forms an angle of less than 45" with the horizontal and a vertical contact that which forms an angle of more than 45" with the horizontal. GRADE DISTRIBUTION: As in rock types, horizontal and vertical grades were defined. ORE HEIGHT: The distance measured between the undercut level and the top of ore. Four categories are considered: (a) less than 60 m; (b) 60 to 100 m; (c) 100 to 150 m; (d)over 150 m

Jan 1, 1981

By Walter Morley

This paper is a record of the first of a series of tests on sulfide minerals to be made by the metallurgical department f the University of California. The purpose of the tests here recorded is to determine the action of standard oils, oil mixtures, and addition agents on the flotation of pyrites. The question of the commercial value of pyrite was not considered. Of the many factors affecting flotation, only two-the oil and the other agent used-were made variable. The flotation agents used are given in the accompanying table. PENSACOLA TAR AND TURPENTINE CO. (PENTARCO) NON-POL., SP. Dis. REF. PER No. KIND GR. POINTS INDEX ?VISC. CENT. 200 Refined wood creosote oil 0.965 105-275 1.5096 1.7 7-8 350 Crude pine wood oil 1.019 70-345 1.525 2.9 4 75 Crude wood turpentine 0.887 65-217 1.456 0.9 10-12 80 Crude pine oil 0.911 70-232 1.4894 1.1 8 750 Heavy pine tar oil 1.063 165-350 1.557 41.0 400 Wood creosote oil 1.025 190-360 1.4977 2.9

Jan 7, 1921

By J. L. McK. Yardley

There are two general classes of problems in connection with electric furnaces. First, those relating to the best utilization of the electrical power inside the furnace; second, those connected with the bringing of the electrical power to the point where it is to be utilized. Referring particularly to the latter class of problems, many calculations have been made, covering a wide variety of furnaces, to check the electrical capacity and power factor found from test. The following analysis, made for the purpose of determining the maximum capacity and approximate performance of a new furnace, to operate at 160 volts on a 60-cycle circuit, may be of interest. The first characteristic necessary to determine is the inductance, since this is the factor that limits the amount of power which can be put into the furnace. Inductance is calculated from the dimensions of the furnace, according to formulas of the U. S. Bureau of Standardsll as follows. Formula 99, for mutual inductance, is: M = 2l [log 2l/d-1 +d/l] Formula 103, for self-inductance, is: Both of these formulas are approximations which are nearly correct when 1 is great compared to d or R. 1 = length of circuit, in centimeters, d = separation between the conductors, in centimeters, R = geometric mean radius of conductors. For a circular tube, R = the radius of the circle. For a solid circular conductor, R = 0.7788 X the radius of the conductor. For a rectangular conductor, R = 0.2235 X (a + b), a being the length, and b being the width of the rectangle. To calculate the inductance of a three-phase circuit, it is customary to figure the inductance of one leg. In circuits where the three conductors

Jan 1, 1920

By Craig R. Barreft, Oleg D. Sherby

The steady-state creep characteristics of pure polycrystalline copper were studied in the temperature range 400" to 950°C and in the stress range 400 to 7000 psi. Tests were conducted in dry deoxidized hydrogen and all specimens exhibited high ductility before fracture. Two activation energies for creep were established as a function of temperature. Above about T/T, = 0.65, the activation energy for creep was 48,000 cal per mole whereas between 0.5 and 0.65 T, it was 28,000 cal per mole. The observed temperature dependence of the activation energy for creep follows the same trend as exhibited by the self-diffusion coefficient of copper in copper single crystals and silver in silver single crystals. It is therefore believed that the rate-cmtrolling mechanism in creep of copper in the stress and temperature range studied is diffusion-dependent and is probably associated with a dislocation-climb process. The low activation-energy region is attributed to enhanced volume difFusion from the presence of dislocations which act as short-circuit paths for diffusion. THERE is extensive experimental evidence1,2 that the activation energy for creep, Q,, for pure poly-crystalline metals tested at temperatures above one half the absolute melting temperature (0.5 Tm) is a constant and equal to the activation energy for volume self-diffusion, Qsd This equality is generally interpreted as meaning that creep in this temperature range is controlled by a dislocation-climb process. The existence of high-temperature creep activation energies not equal to the respective self-diffusion values has been suggested both theoretically3"5 and experimentally.6-8 Two examples of apparent nonequality between Q, and Qsd have been reported for the cases of silver7 and copper,' where activation energies for creep much lower than those for self-diffusion persist to rela- tively high temperatures (up to 0.6 T,). These observations suggest that dislocation climb may not be the only important process in the high-temperature deformation of these metals. That dislocation climb does indeed occur relatively slowly in these two metals due to their low stacking-fault energy has been suggested by many investigators.9-10 Although the creep behavior of copper has been studied extensively at low temperatures, the high-temperature activation energy for creep is not well-defined. The data of Feltham and Meakin8 indicated that Q, is temperature-dependent above one half the melting temperature, equaling a value of -30,000 cal per mole at temperatures up to 550°C and increasing to a value of -49,000 cal per mole at 700°C. No data has been reported above 700°C. The value of 49,000 cal per mole is very near the value of Qsd for copper reported by Kuper et al." as 47,100 cal per mole. In view of the lack of high-temperature creep data and the apparent temperature dependence of Q, for copper the present investigation was undertaken to determine the value of Q, over the temperature range of 400" to 950°C. It was hoped that such a study would clearly document the temperature dependence of Q, and yield significant information as to the nature of the rate-controlling creep mechanisms. MATERIALS AND PROCEDURES The material used in this investigation was high-purity OFHC copper. After rolling and recrystal-lization treatments spectrographic and chemical analyses showed only 0.0008 pct Ag, 0.0002 pctCa, 0.0002 pct H2, and 0.002 pct O2 as trace impurities indicating a purity of greater then 99.995 pct. Specimens were machined from cold-rolled sheet and the gage sections polished with 2/0 emery paper prior to annealing treatments. Annealing was done in both vacuum and dry deoxidized hydrogen atmospheres. No variation in creep strength was found as a result of the different annealing atmospheres. The final annealing temperature was either 700°, 800°, or 1000°C depending on the temperature range to be studied. These treatments resulted in average grain diameters of 0.03, 0.4, and 1.0 mm, respectively. The precise mechanical-thermal history of these

Jan 1, 1964

By Marlin J. Veesaert

Urban quarries offer one of the greatest opportunities for profit in nonmetallic mining with limited financial risk-if properly evaluated, planned and developed to meet market needs. Location, with respect to market, is the prime consideration. Geologic considerations and quality of rock are nearly as important. Cost factors and return on investment (ROI) are the principal financial considerations. Risk of capital is determined by weighing all factors that should be considered and how well judgments are based on information acquired. It is a simple, true axiom of business that an investment must be founded on amount of capital required with its expected return plus profit, expertise required to operate, and time involved. Urban quarries generally offer excellent stone product market opportunities. Costs of land with mineral containment and processing operations are much greater than in rural areas. Environmental problems have become increasingly severe. Competition from other quarries in the market area is usually directly proportional to the profit attractiveness. Risk of capital (initial, developmental and operational) can be evaluated and reduced only by thorough analysis of the entire situation. Time and money expended on planning and preliminary investigations probably represent the most productive investment of available company resources. Basic Reasons to Establish a Limestone Quarry Profit Motive: Certainly the dominant and overriding motive must be that of attaining a reasonable profit. In certain situations a quarry firm may acquire land, necessary rezoning (if required), and satisfy all of the other prerequisites preparatory to beginning operations but delaying the actual development.

Jan 1, 1980

By C. W. Purington, H. B. C. Nitze

The Kotchkar mining-district, known as the Kotchkar System, is situated in the Orenburg government, in Eastern Russia, in the great plateau- or steppe-country immediately adjacent to the eastern slopes of the Ural mountains. Its longer dimension is north and south about 27 miles, while in the widest (northern) portion the distance across is 22 miles. The district lies 40 miles southwest of Miass, a station of the Trans-Siberian railway, and half-way between this village and that of Troitsk. It contains also the not inconsiderable settlements of Kotchkar and Eosobrodskoit., the last a Kossak town. So far as explored, the district is known to contain from 300 to 400 auriferous veins. ? Although the topography is generally of an exceedingly flat character, the country is traversed by several small rivulets running from west to east, and flowing into the Sanarka and Kamenka rivers. Since 1844 placers have been worked in the district, but it was not until 1867 that veins were discovered, and these have been worked only since the impoverishment of the alluvial deposits. The deposit with which this paper especially deals is known as the Ouspensky mine, one of twelve owned by MM. de Zelenkoff & Cie., five of which are situated in the Kotchkar System. An opportunity of visiting this mine was afforded the writers, in eonnection with an excursion of the VII. session of the International Geological Congress; held in Russia during the summer of 1897. The mine is reached by a 40-mile '(troika" journey from the railway station of Bichkil, itself

Jan 1, 1899

By J. Gordon Parr, L. P. Srivastava

DUWEZ1 has shown that pure titanium and pure zirconium transform martensitically during rapid cooling at temperatures about 30° and 15°C re spectively below their To temperatures. Holden et al.2 determined M, temperatures of Ti-Cu alloys by a metallographic technique applied to isothermally treated samples. Their results, superposed on the McQuillan3 version of the partial Ti-Cu diagram are shown in Fig. 1. Sources of metals were: Titanium: Iodide bar; Foote Mineral Co. Ltd. Zirconium: Iodide bar; Foote Mineral Co. LM. Copper: Spectrographic Standard; Johnson Matthey and Co. Ltd. Alloys were prepared by levitation melting under argon to the following nominal compositions:* 0.5, Quenching, at rates to about 8000°C per sec, was performed by cooling in a stream of helium (99.999 pct pure) a 3/64 in. cube of material previously heated by a tungsten coil in a system operating at. 10-5 mm Hg. Cooling rates below 200°C per sec w were obtained by controlling the current to the heater coil. The specimen was supported on a 32 gage Pt, Pt-Rh thermocouple attached through a DC amplifier to an Edin magnetic oscillograph (Model 8082) with a maximum chart speed of 250 mm per sec. Transformation temperatures were read off the oscillograph charts at the location of the beginning of the thermal arrest. The same specimen was quenched several times, and all runs were repeated on new samples. Regular calibrations of the oscillograph against a potentiometer were made between quenching experiments. The criterion of martensite formation was the effect of surface rumpling. Previously polished samples that showed a rumpled surface after transformation are assumed to have undergone martens itic transformation: transformations of a diffusion type produce no surface disturbance, Figs. 2 and 3. Electrochemical measurements were made in a cell described elsewhere4 whose electrodes were the martensitic and the diffusion-transformed materials in the following electrolytes:

Jan 1, 1962

By David B. Brooks, William A. Wallace, James R. Dunn

As the conflict between the mineral industry and preservationists steadily increases, it becomes urgent to determine as precisely as possible the costs of developing vs. not developing our domestic mineral resources so that enlightened mineral policy decisions can be made. A simple discounted cash flow equation is suggested as an approach to assigning a public-interest dollar value to mineral resources. Present value of a piece of land underlain by a mineral resource is taken as equal to the discounted cash flow derived from the resource corrected for present and future surface land values. The equation is effective as a base for mineral aggregate resource planning for metropolitan areas if transportation cost differences are substituted into the cash flow component. Suitably modified, the equation can be fitted readily into an existing computer simulation program for aggregate mineral resource planning in metropolitan areas. However, because of the lack of accurate input, the equation is less effective for other mineral commodities such as base metals and coal. Nevertheless, within a somewhat philosophical context, the equation still has utility and is effective in defining the problem of public interest values for minerals other than mineral aggregates.

Jan 1, 1972

By Joseph Daniels

(San Francisco Meeting, October, 1911.) I. THE FRITZ ENGINEERING LABORATORY. TEE Fritz Engineering Laboratory was built under the direction of John Fritz, and presented by him to the University. A view of the building, looking east, is shown in Fig. 1. . The building was started in 1909, and completed in 1910, although all of the equipment was not placed until later. Mr. Fritz gave his personal attention to the details of construction and equipment, and it was his custom to drive over every day to the University from his home in Bethlehem, and spend some hours watching the work, offering suggestions, making changes, and planning new work. The result is a building and an equipment which embody his practical ideas. The laboratory structure is of the steel mill-building type, of light-colored brick, 91 by 114 ft., of which a section is shown in Fig. 2 and the floor-plan in Fig. 3. The steel frame carries the roof and traveling crane-way. Ample light has been provided by numerous windows in the side and end walls, in the clerestory, and by a skylight, 84 ft. long and 9 ft. wide, in the north roof. The main aisle of the building is 49 ft. 2 in. between centers of crane-columns, and has a clear height of 40 ft. The remainder of the width is taken up by two side aisles, 18 ft. high. The laboratory consists of four sections: (1) A general testing-section, containing the testing-machinery, a small machine shop, and an office; (2) a cement-testing room; (3) a room for making and storing concrete test-specimens; (4) a hydraulic section.

Feb 1, 1912

By E. A. G. Colls

Discusses corrosion in Trail chemical plants producing ammonia, sulphuric, nitric and phosphoric acids, ammonium phosphates, sulphate and nitrate, together with miscellaneous allied material problems and their solution using erosion or corrosion proof materials. Conclusion, that for work of this nature, most costly material may well be cheapest in the long run.

Jan 4, 1950

By Robert H. Hafner, Irvin R. Kramer

IN recent years considerable interest has been shown in surface-alloyed metals, particularly those of chromium (chromized steels), which have excellent corrosion [ ] resistance under a variety of severe conditions, and may undergo bending and flanging operations without spalling. An early method' employed to produce a high-chromium case on steel was developed by Kelley in 1921. It consisted of heating a low-carbon steel in intimate contact with powdered chromium in the presence of hydrogen at temperatures sufficiently high to permit diffusion of chromium into the steel. The usefulness of this process was limited because the extremely high temperatures of 1200° to 1400°C. (2200° to 2550 F.) caused excessive grain growth and warpage. Louenstein and Ulmer2 packed iron castings in ferrochromium and com- [ ] mon salt, and heated for approximately three hours at 930° to 1040°C. (1700° to 1900°F.). A recent and more practical method of depositing chromium on steel surfaces has been developed in Germany by Daeves, Becker and Steinberg3-8 and in Russia by Izgaryshev and Sarkisov.9 This process consists of passing gaseous chromium chloride over the steel to be chromized at temperatures of 900° to 1000°C., or packing them in a porous, ceramic material that previously had been saturated with chromium chloride. They reported obtaining a chromized layer, 0.004 in. thick, on low-carbon steels. Later, Daeves4 established a relationship between the carbon content

Jan 1, 1942

THE book covers principally the assaying of silver and gold and the determination of these values in ores, sweepings, and base metal. It seems to have been written more for the goldsmith and jeweler than for the primary producer of metals. There were approximately sixty different substances needed in the operations described and not over a dozen different types of raw material (ore, matte, crude or alloyed metal) are mentioned. The methods of separation used were essentially simple and reasonably reliable, though the processes were not sufficiently understood to permit the selection of the simplest effective ingredients, and there was a great tendency to employ complicated fluxes, with unnecessary repetition of long-continued heatings indicative of a leisurely pace of life. Practically all the processes described are based on a few chemical facts: the inability of gold to combine with sulphur, while silver and the base metals form easily fusible sulphides; the ready oxidation of base metals; and the ii- solubility of gold in nitric acid. Fluxes serve to make earthy impurities fusible, and to carry oxidizing or sulphurizing ingredients. Many of the processes of assaying had changed little up to the beginning of the present century. Even as late as 1924, when the writer took a University course in fire assaying, the crucibles, cupels, scorifiers, molds, and tongs were identical with those of the sixteenth century, though the wind furnace had a stack on it and the muffle for cupelling was gas-fired.* The fluxes were purer, but of the same essential composition. That this branch- or twig-of metallurgy had not advanced in any funda-

Jan 1, 1949

By G. Rinelli, A. M. Marabini

A wide-ranging series of experiments has been carried out on value minerals (sphalerite, smithsonite and hematite) and gangue minerals (quartz and calcite) to assess the properties of various commercially available vegetable and synthetic tannins as dispersants for these minerals in a finely ground state (<5 microns). Experiments to determine the dispersing power of each reagents on every mineral have been performed through sedimentation tests run first in water and afterwards in the presence of a collector having specific action on the mineral concerned. The tests have been extended to a wide range of dispersant concentrations, maintaining the pH and collector concentrations at the levels known to be the best for the flotation of the various minerals. By comparing the results obtained on each mineral, conclusions have been drawn on the possibility of using different classes of vegetable and synthetic tannins in the selective flotation of fine particles. These results have been related to the structure of the tanning compounds and to the surface properties of the minerals.

Jan 1, 1980

By Robert H. Hafner, Irvin R. Kramer

In recent years considerable interest has been shown in surface-alloyed metals, particnlarly those of chromium (chromized steels), which have excellent corrosion resistance under a variety of severe conditions, and may undergo bending and flanging operations without spalling. An early method1 employed to produce a high-chromium case on steel was developed by Kelley in 1921. It consisted of heating a low-carbon steel in intimate contact with powdered chromium in the presence of hydrogen at temperatures sufficiently high to permit diffusion of chromium into the steel. The usefulness of this process was limited because the extremely high tem- peratures of 1200º to 1400ºC. (2200° to 2550°F.) caused excessive grain growth and warpage. Louenstein and Ulmer2 packed iron castings in ferrochromium and com- mon salt, and heated for approximately three hours at 930º to 1040°C. (1700º to 1900ºF.). A recent and more practical method of depositing chromium on steel surfaces has been developed in Germany by Daeves, Becker and Steinberg3-8 and in Russia by Izgaryshev and Sarkisov.9 This process consists of passing gaseous chromium chloride over the steel to be chromized at temperatures of 900º to 1000ºC., or packing them in a porous, ceramic material that previously had been saturated with chromium chloride. They reported obtaining a Chromized layer, 0.004 in. thick, on low-carbon steels. Later, Daeves4 established a relationship between the carbon content

Jan 1, 1943

By Robert H. Hafner, Irvin R. Kramer

In recent years considerable interest has been shown in surface-alloyed metals, particnlarly those of chromium (chromized steels), which have excellent corrosion resistance under a variety of severe conditions, and may undergo bending and flanging operations without spalling. An early method1 employed to produce a high-chromium case on steel was developed by Kelley in 1921. It consisted of heating a low-carbon steel in intimate contact with powdered chromium in the presence of hydrogen at temperatures sufficiently high to permit diffusion of chromium into the steel. The usefulness of this process was limited because the extremely high tem- peratures of 1200º to 1400ºC. (2200° to 2550°F.) caused excessive grain growth and warpage. Louenstein and Ulmer2 packed iron castings in ferrochromium and com- mon salt, and heated for approximately three hours at 930º to 1040°C. (1700º to 1900ºF.). A recent and more practical method of depositing chromium on steel surfaces has been developed in Germany by Daeves, Becker and Steinberg3-8 and in Russia by Izgaryshev and Sarkisov.9 This process consists of passing gaseous chromium chloride over the steel to be chromized at temperatures of 900º to 1000ºC., or packing them in a porous, ceramic material that previously had been saturated with chromium chloride. They reported obtaining a Chromized layer, 0.004 in. thick, on low-carbon steels. Later, Daeves4 established a relationship between the carbon content

Jan 1, 1943

By Frank C. Roberts

Neglecting the hydrogen and hydrocarbons, 1 will assume the following analysis as a fair average composition, by weight, of the waste gases escaping from a coke-burning blast-furnace: CO2............17.25 CO,............25.25 N,.............57.50 100.00 The 25.25 CO requires for combustion 62.6 of air, or 11.398 of oxygen, necessitating the passage of 48.202 nitrogen into the products of combustion. The latter, therefore, have the following composition: CO2,,............ 56.9 N,............105.7 Tile total specific heats of these quantities of the respective gases are 12.31 and 25.79, giving for the combined product, 38.1 approximately. That is, the product of combustion will absorb 38.1

Jan 1, 1889

By J. R. Cooper, A. Bruynesteyn

The final results of the operation of the Cananea test dumps are discussed. Scale-up factors between shake-flask tests, column tests and the large test dumps are established for the Cananea ore. The formation of clay layers, which formed inside the dumps and markedly affected extraction, as well as secondary enrichment, are discussed.

Jan 1, 1974

By William B. Phillips

The concentration of natural magnetites has been carried on in this country for several years, and more or less information has been collected on the subject. Various inventors, availing themselves of the magnetism which is inherent in such ores, have devised machines for separating the impurities from the ore by means of the electric current; and a greater or less degree of success has rewarded their painstaking and persevering efforts. We are now fairly in position to judge of the value of the process as applied to such material. It is not my purpose to speak of this matter further than may be required by the emergencies of the case. I shall assume that most engineers and metallurgical chemists are already acquainted with what has been done at the Chateaugay mines, at Mineville, at the Benson mines, at the Weldon and Ogdensburgh mines, and at Cranberry; although, at this latter place, there has not been so much done as we could wish. I shall not now speak of the various types of separators. From the time of William Fullarton, to whom was granted an English patent, in 1792, for separating iron-ore by the application of magnetic attraction, down to the present time, a great. many devices have been suggested, and a great many patents taken out for this purpose; but it is not necessary now to enter upon a discussion of their merits. Perhaps, at some future meeting of the Institute, after we shall have experimented further with some of the machines on the market, there may be an opportunity to contribute something to the current knowledge of the efficiency and economy of separating-machines. In this paper I propose to give, as briefly as possible, the results already reached in converting a non-magnetic ore into magnetic ore, and then concentrating it over a magnetic separator of the alternate-polarity type. I am aware of the work

Jan 1, 1896