Search Documents

By G. D. Marshall, D. K. Sarin, V. E. Hampton

Ground support at the Thompson Mine emphasizes scaling, rock bolting and wire mesh screening on the backs and walls of all underground stoping and development areas and also full column grouted cable bolting in the bulk mining areas. This paper describes the equipment for the point anchored rock bolts as well as the full column cable bolts. Modifications in design and testing of all components and procedures has ensured a safe and efficient method of ground support for the mining conditions existing at the Thompson Mine.

Jan 1, 1983

By Alfred H. Bell

Illinois produced 5 per cent more crude oil in 1934 than in 1933, because there was less curtailment, and. the value of the year's production at the wells was 35 per cent more than that of the previous year. The average price of Illinois crude oil in 1934 was $1.13 as compared with $0.87 in 1933, an increase of 30 per cent. The quotas of allowable production for Illinois during 1934, as authorized by Oil Administrator Harold L. Ickes, were as follows: January, February, March, 12,000 bbl. per day; April, May, 12,500; June, July, 12,600; August, 12,500; September, October, November, 12,000; December, 11,700. No other restrictions on production were in effect during the year until Oct. 11, when it was found necessary to prorate the fields to 75 per cent, a curtailment of 25 per cent below the quota fixed by the Federal Administrator. On Dec. 1 the curtailment was reduced to 18 per cent and on Dec. 18, to 15 per cent. Production in Illinois by months during 1934 was as follows: Barrels Barrels January...................... 393,000 July........................ 394,000 February.................... 337,000 August..................... 402,000 March....................... 394,000 September.................. 378,000 April......................... 373,000 October.................... 352,000 May......................... 411,000 November.................. 305,000 June......................... 392,000 December................... 321,000 Total ................... 4,452,000 In spite of higher prices for crude oil, fewer wells were drilled in Illinois in 1934 than in 1933 (Tables 1 and 2). Of the 26 wells drilled, 10 were oil producers, 8 were in proved fields and 2 were in a new pool in Wabash County, l½ miles south of the nearest production, one was a small gas well on the Stubblefield anticline in Bond County, one was an input well for repressuring in the Colmar-Plymouth field, McDonough County, and 14 were dry holes, mostly scattered wildcat tests.

Jan 1, 1935

By AIME AIME

AMERICAN participation in the World Engineering Congress in Japan will be generous. Nearly seventy papers have been prepared and for- warded for printing and the A. I. M. E. is well represented in the list. At our San Francisco meeting four of these will be briefly reviewed by the authors for the benefit of those who are unable to go on to Tokyo. They constitute a group dealing with the present economic status as seen by engineers of the world's iron, coal, petroleum and non-ferrous metals. Messrs. Leith, Smith and Tryon, Requa, and Probert have written papers that constitute a notable series. Messrs. Jeffries, Sweetser and Kinney, and other Institute members have written papers dealing more closely with technologic progress. The latest list of delegates and members is printed below, names of A. I. M. E. members being set in italics.

Jan 1, 1929

By George M. Fowler, J. P. LYDEN

T HE relationship of geological structure to orebodies and to the great masses of chert in the Tri-State mining district is of such significance that it prompts a brief recital of the existing information regarding the various periods of structural deformation. We began to note a relationship between structure and orebodies early in our studies of the field, and further work revealed a like relationship of structure to the great masses of chert so commonly associated with the ore-

Jan 1, 1934

By D. W. Frommer

Processing nonmagnetic taconites by selective flocculation-desliming and flotation requires large volumes of water. If impounded without treatment, these off-process waters require excessively large areas for containment. To discharge the waste water into natural waterways would contribute to stream pollution and likely would not be permitted. In U.S. Bureau of Mines experiments conducted in the Twin Cities Metallurgy Research Center's 900-lb per hr pilot plant, approximately 85% of water requirements for the flotation-based treatment of a Michigan nonmagnetic taconite were met by reclaimed water. Water reclamation of the off-process streams from flotation was accomplished by controlled additions of lime, sodium carbonate, and a synthetic flocculant to reduce turbidities to 51000 ppm equivalent SiO*, while maintaining a Ca(II) content of =16 ppm in the finished effluent. Flotation concentrates of good quality were obtained using the reclaimed water. The cost of chemicals used in water reclamation was approximately equal to the savings in flotation reagents attributed to recycling of the water. Water quality is perhaps as important to flotation as are the reagents used. The character of water is extremely variable, depending on whether the source is a well, lake, or stream, upon the season and temperature, upon prior use, and upon the character of the watershed. All of these factors influence the water hardness and the quantity of other dissolved inorganic salts, turbidity, dissolved and suspended organic matter, dissolved gases, and pH. Frequently, the differences in water quality can measurably influence flotation selectivity, often to the point of spelling success or failure. Water hardness is particularly troublesome in flotation systems employing fatty acids, but other unrecognized constituents may also contribute to peculiarities in flotation behavior. Furthermore, a given water source may be entirely satisfactory in one flotation system, but entirely inappropriate in another. In recent years, society has given more attention than formerly to water use, even in areas where water is plentiful. However, both the demands of a growing population, with increased per capita needs, and also those of industry must be met. As a result, riparian rights must be negotiated with the appropriate government agency or agencies so that consumption of water is often allocated or otherwise controlled. Furthermore, the disposal of off-process industrial and domestic water is coming under the increasing scrutiny of these same governmental units. In these respects, the mineral industry is no exception, so that conservation, water reclamation, and reuse may be expected to assume increasing importance. In 1932, it was stated that, "water reclamation is generally more expensive than the economy in water and reagents resulting from its use. It is employed only if it is urgent to save water."' The economic aspects of this statement may still be true, but the unrestricted use of water is becoming less and less an option of the user. The purpose of this paper is to discuss an investigation conducted by the U.S. Bureau of Mines (USBM), in which water reclamation and treatment were undertaken to develop procedures for, and to assess the effects of, water reuse on the flotation treatment of nonmagnetic taconites. This investigation is a logical extension of previous work described by the author and associates at the Twin Cities Metallurgy Research Center employing selective flocculation-desliming and anionic flotation of silica from low-grade, nonmagnetic iron ores2,8 The requirements for the selective flocculation-de-sliming and the anionic flotation of silica processes determined the direction of the investigation. Previous studies had indicated that pH levels of about 11.0 and 11.8 were required for selective flocculation and flota-tlon, respectively. The calcium content of the water was believed to be of importance in both of these operations, and since Minneapolis tap water with a Ca(I1) content of about 16 ppm had been successfully used, an attempt was made to reclaim the process water at an equal level of dissolved calcium. Additionally, the objective was to nullify or effectively limit the effects of dispersants and fatty acid residuals from prior stages of processing. Last but not least, the system of water treatment had to have the capability of reducing turbidities to workable levels. With about half of the effluent being derived from the selective flocculation-desliming step and carrying about 25,000 ppm of highly dispersed, suspended fines, this last objective appeared formidable at the outset. Procedures for water reclamation were derived, in part, from well-known mineral dressing practices, from past observations and investigations, and from concepts contained in various pertinent publications on water treatment."c These procedures involved: 1) flocculation with lime and poly electrolyte-type flocculants, 2) lime-soda-ash softening, 3) chemical precipitation, and 4) mineral surface adsorption. Control was exercised at various stages of water reclamation by frequent measurements of Ca(11), pH, and turbidity.

Jan 1, 1971

By W. P. Lass

(San Francisco Meeting, October, 1911.) TEE purpose of this article is not only to describe the plant and method of cyaniding the Treadwell concentrates, but to present some of the results of the experimental work obtained in the past three years for the Alaska-Treadwell Gold Mining Co., at Douglas Island, Alaska., under the direction of F. W. Bradley, Consulting Engineer, and Robert A. Kinzie, General Superintendent, of the affiliated companies. At the time the experimental work was undertaken the concentrates were being shipped to the smeltery at Tacoma, Wash., and the cost for treatment of 3-oz. (gold) concentrates was $11.95 per ton, divided as follows: Smelting-charges, . $4.00 Loading, freight, insurance, etc., 2.89 Interest due to time lost in transit and in settlement, . 0.05 Loss due to settlement for 95 per cent. of the gold at $20 per ounce, 5.01 Total, . . $11.95 From the experimental work described later, it was estimated that 96 per cent. extraction could be made by treatment on the ground, and that the cost, when treating 80 tons per day, would be $3.25 per ton, divided as follows; Per Day. Per Ton. Labor, . $66.16 $0.827 Chemicals, . 76.60 0.960 Power and steam-heat, 67.60 0.845 Marketing-, refining- and other charges, 49.36 0.617 Totals, . . $259.72 $3.250 Adding to this total the 4 per cent treatment-loss, which on 3-oz. concentrates amounts to $2.48, gives a total cost of $5.73 per ton. Comparing this with $11.95, the cost when shipping to the smelter, leaves a net gain of $6.22 per ton by the local

Feb 1, 1912

By LEE H. CORNELL

Drilling activity in Kansas during 1945 continued at an accelerated pace to meet the demands of war; 1810 tests were drilled as compared with 1856 tests the previous year, 1771 tests in 1943, and 1516 tests in 1942. During the year 31 new oil pools and 10 new gas pools were discovered. Of a total of 364 wildcats drilled, 323 were dry; and of 1446 tests drilled in pools, 825 were oil wells, 197 gas wells, and 425 dry holes.

Jan 1, 1946

By E. P. Pfleider

Important economic planning by industries, companies, financial firms and governments is predicated on estimates of future growth potential. Prior to the passage of the Taconite Amendment by the people of the State in November 1964, various admittedly conservative forecasts were made of the future of Minnesota's iron ore industry. Since that date, 3 new taconite plants are under construction, 2 are being expanded, one has just gone "on stream," and a seventh is in the design stage. These plants have a total capacity of about 35 million tons of pellets annually. The pessimism of 1964 today is being replaced by unbounded optimism. In the light of such recent developments, it becomes essential to update forecasts related to production, employment and capital investments. New projections show 50 to 58 million tons of pellet production from Minnesota by 1975. This tonnage, together with some 6 to 8 million of natural ores, should provide an annual full-time employment of 15,000 by the mining companies, plus 700 in related railroad and power producing activities and 1000 in plant construction. Total capital investment in new plants should reach $1.5 billion by 1975 and $2 billion by 1990. These current estimates indicate that the conservative goals as set in 1964 for the year 1990 should actually be achieved by 1975 or 1980. BACKGROUND Various forecasts1,2,3 have been made of the future consumption and production of iron ores within the United States. The most recent of these is by Harrison,1 delivered at the American Mining Congress meeting in Las Vegas in October. His estimates reflect the latest iron ore developments on a world wide front as well as the accelerating trend toward pellets. Minnesota's Natural Resources council prepared a detailed report in 1964 outlining the various factors that affect the future of Minnesota's iron ore industry. It presented estimates predicated on the

Jan 1, 1967

By Paul F. Gnirk

INTRODUCTION The notion of permanent disposal of high-level radioactive waste in natural salt formations was formally proposed in 1957 by the Committee on Waste Disposal of the National Academy of Sciences- National Research Council. This proposal was the result of the Committee's deliberations in response to a request from the U.S. Atomic Energy Commission (AEC) in 1955 to assess the viability of geologic formations for such disposal. The research investigations for waste disposal in salt were initiated at the Oak Ridge National Laboratory (ORNL) in the late 1950's and continued through the 1960's and early 1970's at a relatively modest level of funding. During the period of transfer of Federal Administrative authority from the AEC to the Energy Research and Development Administration (ERDA) to the Department of Energy (DOE) during the latter part of the 197O’s, financial support for the research program was enhanced substantially. The program at OWL was regrouped under the Office of Waste Isolation (OWI) and later transferred to the Office of Nuclear Waste Isolation (ONWI) at Battelle Memorial Institute. With this enhancement in the research program, consideration of basalt, granite, shale, and tuff as repository host media was initiated. In February, 1980 the President of the United States established officially a comprehensive waste management program which called for the selection of the first repository site in 1985 from a suite of four or five potentially suitable sites, and initiation of a repository operation at that site in the mid-1990's. The current disposal system concept involves the emplacement of radioactive waste packages in mined disposal rooms at depth in suitable geologic formations, with closure of the facility effected

Jan 1, 1983

By Marius R. Campbell

Does mine-sampling show the commercial value of a coal, and if so, how should it be done? This question is often asked, but seldom answered. During the past summer, while engaged in securing coal for the Government coal-testing plant at St. Louis, I had an excellent opportunity to obtain information on this subject, and the results seem to furnish an answer to the questions propounded above. Sampling for the Geological Survey Coal-Testing Plant. In organizing the work of the testing-plant, it was decided that coal in car-load lots only would be accepted for testing, also that this coal should be loaded under the immediate supervision of a representative of the plant, so as to be certain that the car-load sample would be representative. As an additional check on the work, two mine-samples were to be collected from each mine visited, and these samples, as well as the sam¬ple from the car when it was received at the testing-plant, were to be carefully analyzed at the chemical laboratory, which was established in connection with the plant and operated under the direction of Prof. N. W. Lord, of the Ohio State University, Columbus, Ohio. Mine-Sanmhling:-Mine-samples were collected with great care by me and my assistants, Messrs. J. Shober Burrows, John W. Groves and Frank W. De Wolf. The dominant idea in the regulations which were adopted for mine-sampling was that it should be done in such a manner that the sample would have the same composition as the commercial output of the mine. .

Sep 1, 1905

By J. M. Steininger

The mechanism of flotation of sphalerite with sulthydryl compounds of different acidities has been interpreted in terms of chemisorption of unionized collector molecules on hydroxylated zinc surface sites. At a constant collector concentration of 1.5 x 10 -4 M, the upper pH limit of flotation has been found to be directly and closely related to the negative logarithm of the collector acidity constant, pKa The experimental results indicate that flotation is prevented when about 95% of the collector is ionized. At low pH, collector adsorption decreases because of the reduction in number of hydroxylated zinc surface sites. These two effects explain why sphalerite does not float with the more acidic sulfhydryl compounds. Typical anionic collectors for sulfide minerals are short chain xanthates or dithiophosphates which are nearly completely ionized in the generally alka- line flotation circuits.1,2 Flotation mechanisms have therefore been usually interpreted in terms of ionic collector adsorption and competition with ionic depressants of the same sign as the collector. A few attempts have been made, however, to consider adsorption of neutral collector molecules. Contact angle measurements obtained by Wark and cox3 have been interpreted by Cook and Nixon4 as representing adsorption of the free acid molecules on sulfide minerals such as pyrite, galena, chalcopyrite, and sphalerite. Cook and Nixon tried to show that the free acid concentration of the collector was substantially constant along the curves relating collector addition and minimum pH for bubble adherence. However, in a review of the subject,5 Sutherland presented new data on the flotation of pyrite with mercaptobenzothiazole which did not support the hypothesis of constant concentration of either collector ions or collector free molecules. He further showed that, with the assumption of constant occupation of surface sites, both theories lead to relations between pH and collector addition having similar mathematical expressions. In fact, with the more likely assumption of variable occupation of surface sites which results in more complex relationships, it appears impossible to

Jan 1, 1968

By F. C. Canney

Geochemical prospecting seeks hidden mineral deposits by sampling for variations in the chemical composition of naturally occurring materials. Usually the samples are of soils and other products of weathering and erosion-surface materials extremely susceptible to contamination by human activities. Except when a survey is conducted well away from populated areas, contamination is an ever-present hazard. The program's success is particularly endangered when the contaminants are the elements sought; if these occur erratically throughout an area, spurious anomalies completely unrelated to mineralized areas can be formed. Evenly distributed contamination can be dangerous too, if it raises the background to such a level that true anomalies can no longer be easily detected. Suppose, for ex- ample, that in an area where the lead background is 20 parts per million the soil over a lead vein contains 1000 ppm of lead. Here the contrast would be 50 to 1, which is very satisfactory for geochemical surveying. Now if 1000 ppm of lead from some source were added evenly to the soil in this area, the contrast would be reduced to about 2 to 1, and the anomaly would no longer be readily detectable because a threshold of significance twice the value of the background is the minimum generally used for interpreting geochemical data.

Jan 2, 1959

By Charles T. Baroch, Charles J. Baroch

In 20 years of commercial development, nuclear reactors have demonstrated that they are a safe, dependable, and economical source of power. Operating experience with commercial power reactors has firmly established nuclear power as a practicable source of energy. For both boiling-water (BWR) and pressurized-water (PWR) reactors, performances to date have generally met design predictions. Furthermore, it has become possible to analyze and understand the nuclear, chemical, and physical behavior of this class of reactors. The United Kingdom has had similar success with gas-cooled reactors. The Experimental Boiling Water Reactor (EBWR), the first United States experimental power reactor, began operation in 1956. It proved to be capable of supplying 4500 kw (kilowatts) of power. This reactor was followed in 1957 by the Pressurized Water Reactor at Shippingport, Pa. It was the nation's first full-scale central-station nuclear power plant and the first in the world designed exclusively for the production of civilian electric power. The Shippingport reactor was originally designed to produce 60 megawatts (MW) of electricity, and it was increased to 100 MW in 1964. In 12 years of operation, this reactor generated over 4.5-million MW-hr of electricity in spite of many shutdowns for experimental alterations. By 1969, a total of 15 central-station nuclear power reactors were operable in the United States with an aggregate electrical capacity of over 4200 MW. Early in 1964, Jersey Central Power & Light Co. announced plans for constructing a 515-MW nuclear plant at Oyster Creek, N.J., to be built without government financial assistance. The company stated that a nuclear plant had been selected competitively with fossil fuel. Their economic analysis stated that the plant would generate electricity at less than 4 mills per kw-hr, competitive with coal at 200 per million Btu (British thermal units). Coal cost in 1964 for that region averaged 26.00 per million Btu. The Jersey Central report created wide interest in public-utility circles; nevertheless, no other nuclear plants were announced in 1964. However, in 1965, six nuclear plants were announced by utilities, followed by 26 in 1966, 29 in 1967, and 10 in 1968. By Sept. 30, 1969, 97 reactors were operable, being built, or planned. The cost of these, all committed by electric utilities, was estimated at about $12 billion. When constructed, they will have an electrical capacity of 72,000 MW. The unprecedented surge in power reactors that started in 1965 tapered in 1968, and only seven reactors were announced in

Jan 1, 1976

By H. B. Pulsifier

The following report on the structure of copper is the result of work done in the laboratory of the Rome Wire Co. early in 1925. Previous work had indicated to the author that excellent results might be expected if a suitable technique in surfacing and etching could be developed. The resources of the Rome Wire Co. enabled this metal to be studied more intensively than previously, although none of the materials were investigated exhaustively. The material presented is hardly more than a preliminary summary of the most obvious character of the metal; a much longer research would have been necessary to solve many of the pressing problems connected with the casting and working of copper. The Rome Wire Co. has generously released the paper; to Dr. E. H. Darby, technical supervisor of the company, are due particular thanks for his appreciation of the importance of metallographic study of copper and for his effective support in the prosecution of the work and in making its publication possible. Dendrites, Cells and Granules The microscope reveals at least three orders of structural units in copper, as in other metals. The largest structural units are the dendrites —spine-shaped crystals building up the masses of metal when it solidifies from a melt or is crystallized from solutions by chemicals or electricity. The nucleus cells, commonly called "grains," are the largest and most conspicuous units of recrystallized solid metal; the limitations of their growth usually confine their sizes to microscopic proportions, although in "single crystals" they may become as large as the piece of metal itself. The granules are the microscopic fine-structural blocks or laminations that build up the nucleus cells; they are revealed when metal is etched with suitable chemicals. The granules are commonly visible at 500 diameters magnification and are plainly visible at 1000 diameters; magnification of the granular surfaces up to 6000 to 9000 diameters shows no finer units, not because of lack of optical resolution but because the chemical attack has left an unfeatured smooth surface.

Jan 1, 1926

By J. F. Smith, J. E. Davison

A value for the enthalpy of formation of z2 of -3.14 i 0.21 kcal per g-atom has been measured by the technique of acid solution calorimetry. This result is in quite good agreement with two earlier determinations by tin solution calorimetry and by direct reaction caloriinetry, and averaging of values determined from the three independent calorimetric techniques gives enhanced precision and accuracy with AHh8 (CaMgZ) = - 3.15 i 0.05 kcal per g-atom. For comparison with experimental data, values for the enthalpies of formation of CaMgz, SrMgz, and BaMgz of -9.8, -7.9, and -2.8 kcal per g-atom were estimated from a calculation based on the LVigizer-Seitz approximation as modified by Raimes for polyvalent elements. While complete quantitative accord between these calculated talues and available experimental data is lacking, nonetheless numerical accord is better than might be expected and, more importantly, parallel numerical trends are observed between experimental and calculated vnlues. WITHIN the past decade the enthalpy of formation of CaMg, has been determined a) from measurement of magnesium vapor pressures over binary Ca-Mg alloys,' b) by solution calorimetry with liquid tin as the solvent,' c) from measurement of hydrogen vapor pressures over ternary alloys of calcium, magnesium, and hydrogen,3 and d) by direct reaction alorimetr. The value from tin solution calorimetry is the most precise and is probably the most reliable, and this value is within the quoted uncertainties of the other three experimental results. The overall agreement among the four independent investigations is quite good, particularly so when the diversity of techniques is noted. On the basis of this agreement, CaMgz was chosen as a test material to evaluate the operation of a newly constructed apparatus for the determination of enthalpies of formation of intermetallic phases by acid solution calorimetry. This was believed to be a severe test because of the high chemical reactivity of both calcium and magnesium which reactivity presumably accounts for the fact that an early determination5 of the enthalpies of formation of Ca-Mg alloys by acid solution calorimetry yielded values significantly more negative than the four recent determinations. EXPERIMENTAL APPARATUS AND MATERIALS Experimental Apparatus. The enthalpy of formation of CaMg, was determined by measuring the difference between the heat evolved when dissolving the metallic compound and the heat evolved when dissolving equivalent amounts of unreacted metallic elements in hydro- chloric acid. This was done differentially with an apparatus consisting of twin calorimeters which were constructed to be as nearly identical as possible. The advantage of differential calorimetry is that systematic errors arising from the individual calorimeter design tend to cancel. A schematic representation of the apparatus is shown in Fig. 1. A dead air space around both calorimeters was provided by a large, thermally insulated jacket. Each calorimeter consisted of a 2-liter Dewar flask which was completely enclosed in a copper container. Each Dewar contained 1600 g of 2.5hr HCl to act as the solvent, and thermal effects resulting from solvent evaporation were minimized by covering the acid with 50 g of mineral oil. There was no detectable reaction between the acid and the mineral oil. Equivalent amounts of mechanical energy were added to the calorimeters through twin stirring rods which were driven at the same rpm by a single motor with the intent of the stirring being to maintain thermal equilibrium throughout the solvent. To calibrate the heat capacities of the calorimeters, known amounts of electrical energy could be added by passing measured voltages and currents for known times through submerged heaters, approximately 20 ohms, which were wound noninductively from Manganin wire. A 6-v storage battery was used as a power source, and a dummy heater was used as an exercise circuit to allow the battery to stabilize at a constant electromotive force before energizing one or the other of the calorimetric heaters. A type K-2 potentiometer was used to measure the potential drop across an energized heater while the current was determined from the potential drop across an external standard resistor. Times of energization were measured with an electric timer, and the electrical energy supplied to a heater

Jan 1, 1969

By David White

Jan 1, 1925

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines or phosphate ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pct. When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900" to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By A. Mitchell

An approximate phase diagram has been developed for the CaF2-CaC2 system, indicating a eutectic point at 1240°C, Ncac2 = 0.13, and no detectable solid solution in either phase. The liquidus line is shown to correspond to a simple c22- ion in solution. A thermo-chemical study of&apos; the reaction between carbon-saturated Ni-Ca alloys and CaC2-CaF2 liquids indicates that lhe Raoullian activity coefficient of CaC2 in dilute solution in Cap2 al 1500°C lies between 8 and 10. Some effects of the stabilily of Cap2-CaC2 solutions at high temperatures on electroslag remelting praclice are outlined. THE alkaline earth acetylides. MIIC2, have a reasonably high thermochemical stability at high temperature in the solid state,&apos; with the exception of magnesium, which forms an unstable acetylide at low temperatures (-500°C) and a carbide, Mg2C3, in the range 700&apos; to 1000°C. The acetylides of calcium and barium have been shown to have limited solubility in their respective chlorides,&apos; and further these solutions contain the acetylide as a C: ion.&apos; The equivalent magnesium solutions have not been studied. Although calcium "carbide" is used as a desulfuriz-ing reagent in steelmaking. and is possibly present as an acetylide-oxide phase in very basic electric arc practice slags, the acetylide ion appears to be substantially unstable in a silicate slag.* As a conse- *This instability arises from equilibria in the reaction: CaC2 + CO = (Ca0) + 3C where the low intrinsic solubility of CaC2 in silicate lattice, and the low activity of CaO in a silicate solution where CaO/Si02 < 1, combine to give a very small equilibrium concentration of CaC2 in solution in such silicate slags at temperatures in the region of I 500°c, even under carbon-saturated conditions. Under highly basic conditions, a liquid CaO-CaC2 phase may separate from the silicate system quence of this, the possibility that reactions involving CaC2 in silicate solutions are of importance to general steelmaking practice is remote. However, in operations involving a slag primarily based on a halide, or alkaline earth oxide, we must take into account the possibility that CaC2 will appear in quantities sufficient to significantly affect both the chemical and physical properties of the slag. The work outlined below presents a study of the CaF2-CaC2 system intended to provide sufficient data to allow an estimate of the importance of this system to electroslag remelting and welding practice. However, we should indicate at this point that there will be other processes, e.g., heat treatment, flux cleaning of castings, fused salt electrolysis, and so forth, where alkaline-earth halide fluxes are in contact with carbon, graphite, or carbides, and where halide-acetylide solutions must be taken into account. EXPERIMENTAL 1) Structural Studies. In view of the difficulty ex-perienced in handling CaC2 prepared from calcium turnings and propane gas at 700°C, it was decided to use solutions prepared directly in the equilibration apparatus, Fig. 1. The starting materials were: a) Ni-Ca-C alloy, prepared by adding calcium to liquid nickel held under calcium fluoride in an induction-heated graphite crucible; b) calcium fluoride, prepared by fusing calcium fluoride powder (British Drug House "EXTRA PURE") calcium fluoride in an induction-heated graphite crucible, in air, followed by electrolysis between graphite electrodes at 1 amp cm-2 density, for 10&apos; coulombs per g CaF2. This procedure decomposes the CaO produced by hydrolysis during the fusion step, replacing it by CaC2; Ca2+ + 2e-Ca*(l) Ca*(l) + 2C(gr)-(CaC2)caF2 O2- -2e-O*(g) O*(g) +C(gr)-CO(g) This results in a composition of between 2 and 5 wt pct CaC2 in CaF2. Fifty grams (in lumps) of this material were placed in a graphite crucible, together with Ni-Ca-C alloy (averaging 20 wt pct Ca), and the equilibration apparatus assembled. The alloy reacted with the crucible at high temperature to give CaC2, which dissolved in the calcium fluoride solution to give the desired composition. Cooling curves were plotted manually for these liquids, with rapid stirring and CaF2 seeding to minimize supercooling, and using a Pt/Pt 13 Rh thermocouple calibrated on the freezing points of nickel and copper. This gave a reproducibility of ±0.l°C. and an absolute accuracy of the thermocouple of ±l°C. An example curve is shown in Fig. 2, with the CaF2 end of the binary system in Fig. 3. The CaF2-CaC2 ingots were crushed, under dry nitrogen, and sampled for chemical analysis and X-ray examination. Analytical details are given in the Appendix. Powder diffraction data indicated that the only phases present in all samples examined were calcium fluoride and tetragonal (Types I and 111) calcium acetylide,4 with no evidence of solid solutions or compound formation. 2) Thermochemical Studies. The apparatus used to obtain activity data on CaC2 in these systems is shown in Fig. 4. It consists of an arrangement whereby the graphite crucible and its contents (CaF2-CaC2. Ni-Ca-C) can be rapidly cooled without exposure to air. Trial experiments to determine an equilibration time by ap-

Jan 1, 1969

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines Or phosphate Ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pet&apos; When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900° to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

Jan 1, 1881