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By Donald L. Katz, Leslie J. Noaker

Experimental data are presented for hydrate formation conditions for mixtures of hydrogen sulfide and methane. Vapor-solid equilibrium constants for hydrogen sulfide are also presented. Equilibrium constants, defined here as the ratio of the mole fraction of hydrogen sulfide in the vapor phase (dry basis) to the mole fraction of hydrogen sulfide in the solid phase (dry basis), were calculated using the vapor-solid equilibrium constants for methane presented in an earlier paper. The conditions at which a natural gas will form gas hydrates with water may be computed from its composition, when hydrogen sulfide is present in addition to carbon dioxide and the paraffin hydrocarbons. INTRODUCTION Many gases are known to form solid hydrates in the presence of liquid water at temperatures above 32°F. This hydrate formation is of importance to the natural gas industry inasmuch as most of the constituents found in natural gases will form hydrates. The attention of the industry was called to hydrates by Hammer-schmidtl0 in 1934. A number of papers have been written giving the temperatures and pressures at which these hydrates will form3,4,5,6,7,8,9,10,13,14,15,18,19 and a bibliography on the subject has been compiled13. Deaton and Frost4,5,6,7,8 report work for various pure gases and for some natural gases. Katz and co-workers3,4,13,14,18,19 have reported data for various gases at high pressure. They developed a method of predicting conditions for hydrate formation from equilibrium constants3,19 and from gab gravity." The effects of anti-freeze agents on the formation of hydrogen sulfide hydrate were reported by Bond and Nelson.' The hydrogen sulfide-water system, including hydrate, has been the subject of recent investigation by Sage and co-workers." Mixtures of hydrates behave as solid solutions and an equilibrium exists between the vapor phase and the solid phase that appears to be comparable to the equilibrium between vapor and liquid phases.3,10 By means of the vapor-solid equilibrium constants which have been published for methane, ethane, propane, iso-butane, and carbon dioxidel6 it is possible to determine the temperature and pressure at which hydrates will form for a specific gas mixture when water is present. Hydrogen sulfide which is frequently found in natural gases has a very strong tendency to form hydrate and must enter the solid solution when present in a natural gas forming a hydrate. To determine the effect of hydrogen sulfide when computing the conditions for gas hydrate formation, it is necessary to have the equilibrium constant giving the ratio of the concentration of hydrogen sulfide in the gas phase to the concentration in the solid phase. This paper presents data on hydrate formation for mixtures of hydrogen sulfide and methane and presents the equilibrium constants which were calculated from these data. EXPERIMENTAL PROCEDURE The apparatus used to determine the conditions of hydrate formation for the methane-hydrogen sulfide system was similar to that employed by previous investigators.3,18,19 A charge gas cylinder was used to supply the hydrogen sulfide-methane mixtures, called the charge gas, and to supply the pressure up to 1.000 psia. This

Jan 1, 1955

By D. R. Irving

Prior to this time, ample supplies of high grade mineral fillers, such as clay, have been available close to consuming centers. Now depletion of these accessible deposits, coupled with other factors such as increasing demand due to population expansion and standard of living rise, has made development of upgrading techniques attractive to producers. Various investigations being carried out by the U.S. Bureau of Mines are described. Factors making the development of improved methods for upgrading clays attractive to producers include: depletion of accessible deposits of high-grade clays, rising transportation and labor costs, more exacting specifications, and increased demand resulting from an expanding population and a constantly rising standard of living. Industry recognizes the need for upgrading presently unusable materials and has cooperated with the U.S. Bureau of Mines by providing samples from submarginal deposits and by evaluating the products that USBM upgrades. Following are some examples of recent clay beneficiation work by the Bureau. FIRE CLAY BENEFICIATION At the Rolla Metallurgical Center, Rolla, Mo., the Bureau is conducting research to develop technically and economically feasible mineral-dressing methods to remove quartz, pyrite, and other impurities from submarginal fire clays. The research has been spurred by the growing shortage of high-grade bur-ley, flint, and plastic clays in the Missouri fire-clay district resulting from increased consumption and the low rate of discovery of new deposits of adequate quality. There apparently is ample tonnage of submarginal flint and plastic clays to maintain the industry for many years if such material can be upgraded to commercial quality. Samples have been submitted to USBM by several major producers of refractories, including A. P. Green Fire Brick Co.; Harbison-Walker Refractories Co.; Mexico Refractories Div., Kaiser Aluminum and Chemical Corp.; and Wellsville Fire Brick Co. Research on some samples has been completed and still is in progress on others. The Bureau has found that specimens containing quartz or pyrite in grains coarser than the accompanying clay generally can be upgraded either with a gravity table or hydraulic cyclone. The improvement in quality is notable not only in chemical analysis but in an increase in pyrometric cone equivalent (PCE) of one or more cones. As an example, consider the results attained on a sandy flint—a fine-grained mixture of kaolin and halloysite with quartz and minor quantities of iron and manganese oxides. Chemical analysis was: 35.4 pct alumina (A12O3) and 47.8 pct silica (SiO2). The sample contained 9.2 pct free silica as quartz. The PCE was 33. The sample was crushed to -10-mesh, blunged with water, and then stage-ground to -48-mesh. The ground sample was tabled to yield a clay fraction containing 37.6 pct A12O3, and 45.3 pct SiO2 of which 2.4 pct was free silica. The PCE was 34. Recovery of alumina by this process was 73.6 pct. Concentration of the same clay in the hydraulic cyclone was even more effective in respect to quartz removal. The clay fraction analyzed 38.6 pct A12O3, 45.0 pct SiO2 of which 0.5 pct was free silica as quartz, and had a PCE of 34. Recovery of alumina was 71.9 pct. The hydraulic cyclone was also effective in reducing the pyrite content of a plastic clay which likewise was a mixture of kaolin and halloysite, plus some pyrite, quartz, and carbonaceous material. It analyzed 30.6 pct Al2O3, 52.3 pct SiO2, 2.3 pct Fe2O3, and 0.76 pct S. The sample was blunged and screened through 10-mesh to remove some coarse pyrite, then ground to -65-mesh and cycloned to remove most of the remainder. The final product recovered 81.8 pct of the alumina at a grade of 32.3 pct A12O3, 50.9 pct SiO2, 1.4 pct Fe2O3, and 0.1 pct S. The PCE was 32-1/2, compared with a PCE of 31-1/2 for the crude clay. The fine beneficiated clays, of course, must be pelletized and dried or fired to meet commercial acceptance. This operation has already been solved by industry. Little progress has been made on two remaining problems. One is that of fine-grained impurities, not separable by table or cyclone. Flotation is being investigated to solve this. Although fine pyrite is readily rejected, a satisfactory quartz-clay separation has not been attained. The second problem, a stubborn one, is that of alkali. Samples containing as much as 5 pct potas-

Jan 1, 1961

By H. E. Montgomery

Engineering trends for some years have been toward higher temperatures, both in process work and in steam generation. Progress in design has been dependent upon the development and evaluation of suitable alloys. Elementary knowledge of the limitations of copper-base alloys indicate their unsuitability for the higher ranges of operating temperatures; say, for steam to 950°F. Emphasis upon the more severe conditions perhaps is a reason why more complete evaluations have not been made for copper-base alloys over their range of suitability. However, the few studies made show clearly the need for more accurate evaluations. Some copper-base alloys long looked upon favorably because of superior room-temperature properties are not as desirable as some others perhaps not so "strong" and "hard," as measured by usual acceptance tests. Load-carrying capacity and structural stability arc two important factors. Structural stability as influenced by temperature, by presence of corrosive media, or by the effects of stress itself, frequently is a limiting factor. An early reference to creep properties of the cast copper-base alloys is that of Brown1 (Admiralty gun metal, Cu 87.5, Sn 10.1, Zn 2.3 and Pb 0.08). His tests were similar to present-day stress-rupture tests without measurement of creep elongrttion. Loading was in steps with usually 24 hr. between steps, and continued until failure. He advocated a "useful strength value" of 61/2 tons (I4,560 Ib.) per sq. in. at 572°F. (300°C.). A specimen broken in 1/2 hr. at 572°F. had a strength of 21 tons (26,880 Ib.) per sq. in. The strength values as determined by 1/2-hr. tests and those by the longer stress-rupturc start to differentiate at 392°F. (200°C). This is the temperature at which time at load began to play an important part. Bolton, Price and Hebemann2 give creep curves on 88-10-2 (A.S.T.M. B-60, CU 88, Sn 10, Zn 2) and steam or valve bronze (A.S.T.M. B-6I, Cu 87.50, Sn 6.00, Zn 3.75, Pb 1.70) at 500°F. Direct comparison of the two alloys under 7500 lb. per sq. in. stress showed the steam bronze (A.S.T.M. B-61) superior in creep strength to the 88-10-2 (A.S.T.M. B-60). The steam bronze shows no creep after 700 hr. (1540-hr. test). The 88-10-2 shows a rate of about 7.0 per cent in 100,000 hr. over the last 700 hr. of the test. Creep curves on the steam bronze are also given at the same temperature (500°F.) with loads of 6500 lb. per sq. in. and 8000 lb. per sq. in. These indicate "creep strengths" (creep rate within I per cent in 100,000 hr.) in excess of 8000 lb. per sq, in. for this alloy. It was found that higher percentages of tin are detrimental to creep resistance. This is caused by the presence of or formation during test of considerable alpha-delta eutectoid in the higher tin alloys. Bolton also states that the high

Jan 1, 1945

By G. L. Craig

COPPER containing a small amount of arsenic has been used extensively abroad, and a number of reports, particularly those by English workers, have indicated that such copper is better suited for certain uses than high-conductivity copper. Inasmuch as a large percentage of the copper fabricated is used for nonelectrical purposes where high conductivity is not at all essential, it has been suggested that arsenical copper might be at least as suitable as high-conductivity copper for many purposes. An investigation spon-sored by the Calumet & Hecla Consolidated Copper Co. at Battelle Memorial Institute was therefore undertaken with the object of determin-ing the effect of small amounts of arsenic (up to 0.75 per cent) on the properties of copper. In this work the effect of arsenic on the basic properties of tough-pitch and of phosphorized copper have been critically studied and many data collected from laboratory experiments which have been supplemented by mill-scale tests. The results of these tests, with coppers of the type referred to, in the annealed condition and after various degrees of reduction, show that the basic properties of copper are not greatly affected by small amounts of arsenic, an increase in strength being observed with increasing arsenic with no corresponding decrease in ductility. It is well known that the evaluation of material for plastic deforma-tion, on the basis of laboratory tests, may not be directly comparable with its performance in commercial operations. In order, therefore, to substantiate the data obtained at the Institute, the cooperation of the Wolverine Tube Co. was obtained for a series of production tests on the fabrication characteristics of phosphorized arsenical copper in the manu-facture of seamless copper tubing. In these tests the behavior of the arsenical coppers was carefully observed in comparison with that of representative types of phosphorized arsenic-free copper, which were fabricated under the same conditions.

Jan 1, 1933

By F. L. Wolf

THE reclaiming of metallics from slag and sweepings is of vital interest to every brass-foundry man, but the first cost and interest on the investment often make it prohibitive for the small foundry to enjoy the benefits of a complete concentrating unit. However, the cost of such an installation and the cost of operation are often exaggerated to such an extent that foundries dispose of their copper-bearing refuse at a great loss. This paper gives the actual costs and the returns obtained in the reclaiming plant used at the Ohio Brass Co. This plant is divided into two sections, the preliminary treatment and concentrating departments. In the former occur the screening, crushing, and hand-picking operations whereby most of the non-valuable and heavy metallic constituents are removed. The concentrating equipment consists of an elevator, bins, jigs, ball-mill, and table. PRELIMINARY TREATMENT The preliminary treatment varies, depending on whether the material is slag, foundry-floor sweepings, or brass casting-floor sweepings. Material of the first class is produced from the operation of six Steele-Harvey and six Schwartz furnaces; the second class represents sweepings from an average of thirty-five molding floors per day; and the last is collected around the sprue cutters. and knockout bench.

Jan 9, 1920

By Alfred F. Beasley

The slags derived from the smelting of lead and copper ores are composed essentially of silicates. The problems arising from the smelting of these ores consequently involve the study of silicate fusions. In the absence of specific knowledge concerning the compounds occurring in slags, it has been the practice of metallurgists to treat them in terms of their total chemical composition. Slags have been classified as monosilicates, bisilicates, sesquisilicates, and subsilicates according to the relative amounts of oxygen in the acid and basic radicles of the hypothetical salt produced. While this has been a useful guide to metallurgists in computing the furnace charges, it falls far short of a basis for the scientific investigation of slags. In order to understand the nature of a slag as a whole it is necessary to learn the physical and chemical properties of the mineral compounds which go to make up that slag. The efficiency with which an ore is smelted in a furnace depends primarily on the physical properties of the compounds formed during the operation. When the melting points and fields of stability of each of the mineral compounds in slags have been established, it will be possible to determine the properties of any given slag by examining the nature and amounts of its mineral constituents. Earlier Investigations In 1884 and succeeding years an extensive survey of lead and copper furnace slags was made in Europe by Vogt, and a large part of the present knowledge of the subject is due to his publications.' Vogt's interest in artificial fusions was directed primarily towards the explanation of problems in igneous and economic geology and his subsequent works have been confined wholly to that subject.

Jan 1, 1930

By Orson Shepard

THE leaching method that was first widely used with the cyanide process consisted of percolation leaching of crushed ore in vats or leaching tanks. It was frequently necessary to separate the sand for cyanide treatment and discard the slimes, because only granular material allows the necessary amount of solution to percolate through it in a reasonable time. Later a slime process consisting of agitation followed by thick-ening and filtration was developed, and is extensively used to the pres-ent day. Even though the agitation-leaching slime process is considered satis-factory, the cost of plant and of operation are both greater than for a percolation-leaching plant of equal capacity. This situation has been recognized in the hydrometallurgical treatment of copper-bearing slimes; and a considerable amount of j work has been done at the Southwest Experiment Station of the U. S. Bureau of Mines in an effort to develop a percolation method of treating slimes. Experimental work on the copper-leaching problem was started by H. E. Keyes1, and later tests were made by J. D. Sullivan and A. P. Towne2. The Bureau of Mines work resulted in the development of a process by which slimes were moisture-agglomerated with sand and crushed ore, to form a permeable mass through which solution could be trickled without breaking clown the glomerules. The agglomeration trickle-leaching method was devel-oped on a small scale and finally tried out in a 12-ton vat. Successful results were obtained as long is the leach solution was added slowly, so that the agglomerated charge was never flooded. If the leach solu-tion flooded the charge, the glomerules broke down and, the charge became impervious. There are two important difficulties in the application of the agglom-eration trickle-leaching-method to the cyanide process for gold and silver ores:

Jan 1, 1937

By D. R. George, J. R. Ross, J. D. Prater

In the United States nearly 200,000 tons of copper per year are being produced, by leaching waste rock and oxidized copper ore with dilute sulfuric acid-ferric sulfate solutions and precipitating the copper on scrap iron. The flow of solutions is in excess of 80 million gpd. A survey conducted by the U.S. Bureau of Mines in late 1965 at 14 mines in Arizona, Utah and Nevada, showed that with a few exceptions, the solutions contain from two to a maximum of 15 ppm U3O8, and there is strong evidence that these represent equilibrium concentration and not merely an accumulated in-process inventory resulting from recycling the solutions for many years. Consequently, it is believed that uranium could be recovered at a constant rate as long as mining and leaching operations continue. Preliminary surveys and estimates by the USBM indicate a production potential of possibly 1000 tons of U3O8 per year at a cost of less than $8 per lb if the major process streams from most of the large copper mines in the western United States were to be treated. Laboratory research by the USBM showed that the uranium could be efficiently recovered as a specification-grade uranium concentrate by a combination of ion exchange resin and solvent extraction of liquid ion exchange techniques. A joint pilot plant investigation by Kennecott Copper Corp. and the USBM was conducted on solutions from the Bingham Canyon mine to evaluate a new system of countercurrent ion exchange developed during the laboratory studies. In subsequent small-scale continuous tests, recovery of uranium from the pregnant eluate by solvent extraction and chemical precipitation techniques was investigated.

Jan 1, 1968

By Henry M. Howe

My attention has been drawn within a few days to a series of articles in Volume XVIII of the Engineering and Mining Journal, 1874, by Mr. J. A. Church, in which it is stated, among other

By J. M. W. Mackenzie

Recent successful applications of emulsion flotation suggested that a detailed study of some of the physicochemical aspects of the process was warranted. The objective of the current research was to provide general information on the electrokinetic properties of emulsion drops of the type commonly used in oxide mineral flotation. Combinations of the parafin oil Nujol and several long-chain reagents in water were studied with respect to the effect of pH on the surface and zeta potentials. It war concluded that for emulsion drops in the presence of a long-chain reagent but in the absence of any other electrolyte which is adsorbed at the oil-water interface, the surface potential will be some two to four times the value of the zeta potential. During the last decade there has been a revival of interest in the so-called agglomeration, or emulsion flotation (in this paper the term emulsion flotation will be used), process. In this flotation technique, a neutral hydrocarbon oil, together with longcnain reagents, usually of a polar-nonpolar structure, are added to the pulp to function as collectors for specific minerals. The combination of oil and longchain reagent is either emulsified prior to addition to the pulp, or in the pulp itself, and the resulting collector species is an oil-in-water emulsion of a neutral hydrocarbon oil emulsified by polar-nonpolar long-chain reagents. Emulsion flotation has been used successfully in the treatment of ores containing iron, manganese,' and titanium oxides. 2 Two other major and long established applications of emulsion flotation are the flotation of coal and Florida phosphate ores. In recent years, emulsion flotation has been particularly successful in the treatment of ores containing oxide minerals which slime readily and, consequently, are difficult to treat by conventional flotation processes. The recent successful

Jan 1, 1970

By Gilbert Thiessen

The use of a pneumatic process for removing fine material (usually less than 48-mesh) is increasing in the preparation of stoker-size coal. The fine material, produced as a byproduct, is known as dedusting-plant dust. Since fusain is the most friable ingredient in bituminous coal (it may readily be rubbed to an impalpable powder with the fingers), it tends to accumulate in this byproduct. While studying the properties of cokes produced by blending Illinois coals with other materials, considerable quantities of fusain were required. Dedusting-plant dust suggested itself as a convenient supply. It became necessary, therefore, to determine how much fusain was contained in a measured sample of dust, how the fusain was distributed with respect to size and specific gravity in the sample, and whether or not it would be feasible or desirable further to concentrate or to recover the fusain from the dust before using it in the blends to be coked. The results obtained indicate the possibilities of production of a high-fusain material. Previous Work Lehmann and Stach have developed a processl,2 for separation of banded bituminous coals into their band ingredients by impact pulverization followed by screening or air separation. They found, as might be expected, that the fusain was concentrated in the very fine material. Prockat and Raute3 examined samples from dedusting plants operating in the Ruhr. They separated the dust into various screen sizes and determined the composition of each of these sizes in terms of the banded ingredients of the original coal. In the dusts examined by them, the finest material was always high in fusain, and most of the fusain contained in the original samples could be recovered from this fraction.

Jan 1, 1936

By Gilbert Thiessen

The use of a pneumatic process for removing fine material (usually less than 48-mesh) is increasing in the preparation of stoker-size coal. The fine material, produced as a byproduct, is known as dedusting-plant dust. Since fusain is the most friable ingredient in bituminous coal (it may readily be rubbed to an impalpable powder with the fingers), it tends to accumulate in this byproduct. While studying the properties of cokes produced by blending Illinois coals with other materials, considerable quantities of fusain were required. Dedusting-plant dust suggested itself as a convenient supply. It became necessary, therefore, to determine how much fusain was contained in a measured sample of dust, how the fusain was distributed with respect to size and specific gravity in the sample, and whether or not it would be feasible or desirable further to concentrate or to recover the fusain from the dust before using it in the blends to be coked. The results obtained indicate the possibilities of production of a high-fusain material. Previous Work Lehmann and Stach have developed a processl,2 for separation of banded bituminous coals into their band ingredients by impact pulverization followed by screening or air separation. They found, as might be expected, that the fusain was concentrated in the very fine material. Prockat and Raute3 examined samples from dedusting plants operating in the Ruhr. They separated the dust into various screen sizes and determined the composition of each of these sizes in terms of the banded ingredients of the original coal. In the dusts examined by them, the finest material was always high in fusain, and most of the fusain contained in the original samples could be recovered from this fraction.

Jan 1, 1936

By E. J. Duggan

It is a coincidence that the two plants in this country for treating copper ores by ammonia leaching came into existence at almost the same time, that of the Kennecott Copper Corporation at Kennecott, Alaska, in March, 1916, and that of the Calumet & Hecla Mining Co. in July of the same year. The success of both from their very beginning attests the merits of the process under widely different sets of conditions. As would be expected, in the time they have been operating these two pioneers in this field have greatly perfected their methods and improved their technic. The early difficulties inevitably encountered in undertakings of this nature have been largely surmounted and the process may now be considered to be well seasoned and matured. There have been no other commercial plants built in this country and but one elsewhere, the Bwana M'Kubwa in Northern Rhodesia, for the reason that this method of treatment may be employed on comparatively few ores. However, it has certain advantages, some of which are inherent to sand leaching processes generally and some to ammonia leaching in particular. Among these may be enumerated: (1) the avoidance of fine grinding; (2) the small amount of water required; (4) a high-grade product; (5) little likelihood of deleterious gangue constituents and fouling of solutions. It is true that great progress has been made in recent years in the flotation of the ores that come within the province of ammonia leaching. Nevertheless ammonia leaching has its possibilities and these may be overlooked in the present tendency toward the use of flotation for all ores. Probably the chief requisite to its further extension to suitable ores will be a cheap source of steam. Ammonia, the only reagent required, is unlimited in supply, readily transported and the price that is likely to obtain in the future should not make its cost at all onerous. Choice of Method for Kennecott Ores The Kennecott ores are typical limestone replacement deposits. The milling ores have a copper content of approximately 8 per cent in a limestone dolomite gangue. About half the copper occurs as sulfide, mostly chalcocite with some covellite, and the remainder in the form of the two carbonates, malachite and azurite. A comparatively coarse grind followed

Jan 1, 1934

By E. Gybbon Spilsbury

It is not my purpose in this paper to describe all the different contrivances which have been devised, during the last quarter of a century, for the purpose of Iessening the expense of mining and tunnelling operations; I wish merely to call your attention to the crying want of the Western miners for some cheap and effective machine, capable of economizing on the heavy item of hand-drilling, and to enumerate a few of the latest improvements which have been made in machines of this class. This necessity for the introduction of some kind of labor-saving machinery in mining, especially in the West, begins to make itself felt more imperatively every day. No branch of industry has remained so conservative to the old customs and systems of working, as has mining; and yet no branch is so entirely dependent on economical working for a successful issue; and in very few is the proportion of cost of hand-labor to profits obtained so preponderantly heavy. Perhaps the chief reason for this seeming apathy towards progress in this direction arises from the fact that nearly all mining operations, especially in this country, are carried on in regions more or less remote from the great centres of civilization, thereby rendering the carriage and erection of machinery extremely difficult and expensive. Another reason, however, exists in the fact that a large number of men in charge of mining operations are greatly prejudiced against the introduction of machinery underground, from the idea that it cannot be made to adapt itself to the various positions and

By James R. Garvey

Coal reserves of the United States are enormous. Considered on the basis of proven reserves, and compared with reserves of other fuels, coal constitutes 88% of the proven recoverable energy resources of the nation. That includes oil, natural gas, oil shale and uranium used in burner reactors. Considered on the basis of "ultimate recoverable reserves", that is, those resources which can be extracted with improved technology and at somewhat higher costs and including those resources of all fuels not yet found but believed to be waiting discovery, coal still accounts for 75 % of the national total. In the U.S. hydrocarbon economy about 76% of the total energy needs are supplied by petroleum and natural gas and about 20% by coal. Domestic resources of oil and gas are running dangerously low, and assuming a continued hydrocarbon economy, the 20% supplied by coal may increase dramatically. Nearly a quarter of the petroleum consumed in the U.S. is imported. Natural gas is in shorter supply than petroleum and the problems of its importation seem at least as difficult as those of oil.

Jan 6, 1972

COMPARED with the equipment used in most industrial processes, the open-hearth furnace has a relatively short life. The most important quality of an open-hearth refractory, therefore, is its rate of failure. This discussion of the physics and chemistry of the individual refractory materials will keep uppermost the question "Why, how, and where does it fail?" CAUSES OF FAILURE The principal causes of failure of open-hearth refractories are, in their order of importance : 1. Reaction with iron oxides. 2. Reaction with slag. 3. Flow and failure under load. 4. Melting. 5. Spalling. 6. Mechanical breakage. Reaction with Iron Oxides and Slag. Reaction with iron oxides and slag is by far the most important cause of failure. Liquid iron oxide comes near to being the "universal solvent" sought for by the alchemists. Among the materials commonly used in the open-hearth furnace, magnesia has the greatest resistance to this liquid and fireclay the least. The relations of iron and oxygen at high temperatures are illustrated in Fig 14-18. The liquid iron oxide that forms wherever iron is in contact with air or combustion gases cannot be given a definite formula. The percentage of oxygen at steelmaking temperatures may range from 21 to 28 (iron 79 to 72), depending upon the temperature and the oxygen pressure and also upon whether or not the liquid is in equilibrium with metallic iron. Oxide mixtures exist not only in the solid and liquid phases but also in the vapor phase, for metallic iron and its oxides have

Jan 1, 1964

COMPARED with the equipment used in most industrial processes, the open-hearth furnace has a relatively short life. The most important quality of an open-hearth refractory, therefore, is its rate of failure. This discussion of the physics and chemistry of the individual refractory materials will keep upper- most the question "Why, how, and where does it fail?" CAUSES OF FAILURE The principal causes of failure of open-hearth refractories are, in their order of importance: 1.Reaction with iron oxides. 2. Reaction with slag. 3. Flow and failure under load. 4. Melting. 5. Spalling. 6. Mechanical breakage. Reaction with Iron Oxides and Slag. Reaction with iron oxides and slag is by far the most important cause of failure. Liquid iron oxide comes near to being the "universal solvent" sought for by the alchemists. Among the materials commonly used in the open-hearth furnace, magnesia has the greatest resistance to this liquid and fireclay the least. The relations of iron and oxygen at high temperatures are illustrated in Fig 14-18. The liquid iron oxide that forms wherever iron is in contact with air or combustion gases cannot be given a definite formula. The percentage of oxygen at steelmaking temperatures may range from 21 to 28 (iron 79 to 72), depending upon the temperature and the oxygen pressure and also upon whether or not the liquid is in equilibrium with metallic iron. Oxide mixtures exist not only in the solid and liquid phases but also in the vapor phase, for metallic iron and its oxides have

Jan 1, 1951

By Mack H. Shumate

1 Raw coal from Truax-Traer Coal Co.'s Burning Star Slope mine, Jackson County, Ill., is brought to the surface by the new 42-in. slope belt haulage conveyor at the rate of 850 tph. The loading system shown above employs four storage tracks, two tail tracks for loading, and four 25-hp hoists, operating in pairs, and coupled to two dolly cars to move the trains of cars through the tipple. The system has a capacity for loading 1800 tph. One operator is needed to load on each of two shifts. A third man working for a few hours on night shift sets up for the morning cycle. Railroad crews bring in empties placing them in the storage yard, twenty cars to the track. Trains of empties are pulled through to the tail track and loaded on the return; hoists at opposite ends pull the trains in each direction. [ ] 2 Pairs of hoist drums are interconnected electric- ally by a Trench lay cable buried between the tracks. When one hoist operates moving the cars, the tail hoist idles with brake released. The operator, situated in the tipple over the cars, views the storage yard at the end of the first shift (left). When returning loaded cars to yard, four or five cars are held on the dolly to couple the next string of empties. Empties are ready for loading on the tail track (right). Differing from a gravity system in which empty and load tracks are on a slight grade, this system is designed for 80 railroad cars for a two-shift operation and one rail service per day.

Jan 6, 1962

AT THE beginning of September, the French eased their pressure against Germany and allowed the latter to, be granted an arrangement for its reparation payments in kind according to the Lubersac-Stinnes agreement. The whole arrangement as to the German reparations is, however, but temporary. The Greeks suffered a disastrous defeat at the hands of the Turks in Anatolia, which rapidly resulted in the expulsion of the Greeks from Asia Minor, the Turks occupying Smyrna about Sept. 10. A few days after entering Smyrna the Turks sacked the city and finally set it afire, creating one of the most terrible destructions in history. At the same time their advance on Constantinople, together with their demands on the Allies, brought them directly in, con-flict with the latter and Great. Britain prepared for a new war, with doubt as to the support of France, whose political interests in the Near East are as different as their economic interests with respect to ;Germany. The Turkish situation became thus a threat against the peace of Europe. A revolution in Greece, followed by the abdication of King Constantine and the return of Venizelos to a function in the Government was one of the conse-quences of the defeat in Asia Minor. Civil war continued in Ireland, with more or less -bloodshed. Economically there has been no change for the better in Europe. In Great Britain there, are no signs of an -important trade revival. In Germany and Austria things are worse. French, Italian and Spanish affairs continue in a bad way. In the United States business is represented as "booming," but many people are inquiring where is the boom? About the middle of September Congress passed the tariff and the soldier bonus bills. The President signed the former and vetoed the latter, with a strong message that won praise among people who understand condi-tions. Congress tried to pass the bill over the veto, but failed. The President signed the tariff bill, which became effective Sept. 21. A severe handicap was thus imposed on American economics.

Jan 10, 1922

By Albert E. Reed

THE development of the modern high-speed vibrating screen, together with the increasing availability of long-lasting stainless-steel screen cloth for relatively fine-mesh separations, means that more screen circuits will be used in the future. It seems pertinent therefore to examine a few characteristics of screen-circuit products, and especially the points of difference as compared with products of classifier circuits. The most noticeable change that takes place when substituting screening for classification in a grinding circuit is a substantial drop in the circulating load, owing to the greater ease with which the "fines" leave the circuit. Where circulating loads of 500 to 600 per cent are considered fairly normal in classifier grinding circuits, the grinding circuit equipped with screens for equivalent work will have a circulating load of only 20 to 75 per cent. As a result, the products of screen circuits will show a minimum of overgrinding or sliming as compared with a classifier circuit. The desirability of this characteristic will naturally vary according to the metallurgical requirement in each case. Because screen separations make a sharper division according to particle size, the products of screen circuits can be expected to show greater crowding of particles near the point of separation. These changes in product characteristics are so important that other factors, both in the grinding and concentration processes must be adjusted if the operator is to take full advantage of the increased tonnages and higher efficiencies available to him through the use of the screen circuit. One factor that too often is overlooked when changing to screen circuits is the importance of adjusting the ball-mill or rod-mill charge, to compensate for: (I) increased tonnage of new feed, (2) decreased circulating load, or (3) more granular character of feed. These same factors should also be taken into consideration for their effect on the various types of concentration or flotation equipment that may follow the screen circuit. Over a period of years, this equipment has generally been designed and adjusted to the product of the classifier, therefore it seems obvious that some research on the adjustment and adaptation of concentrating and flotation equipment to screen-circuit products will be most profitable to the operators. To illustrate the points mentioned above, the results from three recent installations follow, with data on the comparative characteristics of screen and classifier circuits. A range of separation from 10 to 48 mesh has been chosen, using two general types of screening equipment and two types of classifiers. The first of the installations of which the results are to be examined is that of an iron-ore installation where eastern magnetite was being concentrated prior to sintering. Fig. I shows the original grinding circuit, using classification, while Fig. 2 shows the same circuit equipped with a

Jan 1, 1945