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By Dever Ashmead

ANTHRACITE was first mined in the Wyoming Valley and sold as an article of commerce in 1808. As some preparation has always been necessary to make it ready to burn, the preparation of anthracite must date back over a hundred years. Two vital factors have determined to a large extent the degree and method of anthracite preparation. These are, first, the character of the beds worked and the methods by which they were mined and, second, the equipment used and practices followed in the burning of the coal. It is not the intention here to go deeply into these phases of preparation as a paper of no mean length could be prepared on the history of either. Rather, the intention here is to point out the main considerations and to show the influence they have exerted on the preparation of anthracite. MINING METHODS AND THEIR RELATION TO PREPARATION In the beginning f the nineteenth century, the coal beds were virgin with the possible exception of some outcroppings that had been worked, to a slight extent, by the Indians. It is known that the American aborigines had a knowledge of the use of this fuel, because when the Wyoming Valley was purchased from them, in 1754, they mentioned the fact that, by selling the land, they would lose their coal.

Jan 9, 1921

By CHRIS A. HARTNAGEL

For the second consecutive year, the production of crude petroleum in New York has fallen below the 5,000,000-bbl. Mark that had prevailed previously since 1937. In 1945, the output totaled 4,658,000 bbl., a decrease of slightly more, than one per cent from the 1944 production. This total represents 2 per cent of the stripper-well production of the United States, which in turn is I 3 per cent of the total annual yield of the nation's petroleum.

Jan 1, 1946

By J. H. East, E. R. Maize

Probably no other type of equipment is now being introduced into American underground mines about which less is known and about which there is more misinformation than the Diesel mine locomotive. Any equipment installed underground brings with it hazards that were not present before, and it is the purpose of this paper to show that Diesel-powered equipment can be used with safety in underground mining if certain feasible but very important precautions are taken. Diesel locomotives are not new in underground mining, since they have been used successfully in European mines for at least 19 years, but their use in American mines is exceedingly limited; so far as is known, only five Diesel mine locomotives are operating at present in United States mines. Diesel locomotives were permitted underground in Great Britain for the first time in 1939, and the rapid increase in the number in use in those mines testifies to their safety and efficiency. More than 1300 Diesel mine locomotives are in use in European mines, according to the latest statistics; these locomotives are employed in coal, iron, lead-zinc and mercury mines. SO far as is known, the Diesel mine locomotive has not yet caused an explosion or mine fire, and has not been the cause of any fatalities from toxic gases in the mines in which it has been used. Adjustments of Diesel Engines foR Underground Use A Diesel engine adjusted to work in an underground mine should not be confused with the same engine adjusted for use outside the mine. Adjustment of the air:fuel ratio for engines that are used on the surface is intended to give the maximum fuel economy, and little or no attention is given to the chemical composition of the exhaust gas produced. Many such engines are poorly adjusted and perhaps poorly maintained, with the result that the engine emits smoke in quantity as well as the disagreeable odor familiarly associated with Diesel engines. The air:fuel ratio in the case of Diesel engines operated on the surface is generally around 15:1; that is, 15 lb of air to burn I lb of fuel. Diesel engines adjusted to work underground usually have an air:fuel ratio of about 20:1, as at this ratio the minimum quantity of toxic gases is given off in the engine exhaust. Diesel engines operated underground must be maintained in better condition than similar engines used on the surface; otherwise, the smoke emitted in the exhaust becomes objectionable if not unendurable. Practically normal air is required for safe operation of Diesel engines underground or in confined places; failure to provide virtually normal air for Diesel engine operation is almost certain to result in trouble. Bureau of Mines engineers have determined that the air in which a Dicsel Operates hi confined places should contain less than % pct of carbon dioxide and at least 20 pct of

Jan 1, 1949

By B. M. Larsen

A discussion based on three commercial furnace tests and electrical analogue calculations is presented. It shows that while regenerator efficiency is mainly dependent on loading or relative amount of heat exchange surface plus the heat transfer coefficients as effected by flow velocity, the actual air preheat temperature can be high or low, largely independent of regenerator efficiency, mainly due to the diluting or unbalancing effects of cold air leakage into various parts of the furnace system. THERE are many signs of a renewal of interest in the heat regeneration aspect of the open hearth process, among them being the recent paper by Marsh,' the introduction of auxiliary checkers in the stacks of the Isley furnace design, and a number of recent installations of two-pass checkers in new and old furnaces. However, there are also signs of yielding to the usual temptation of oversimplifying a problem. Faced with the job of obtaining better draft on a furnace, for example, a good designer always recognizes that focusing attention on some one factor, such as height of the stack or capacity of the exhaust fan, does not make his problem any easier; rather he must look at the whole system, including such items as valve resistances, bends in the flues, and flow resistance in the waste heat boiler; then he must make his design in relation to the available draft so as to avoid any serious bottleneck in the system. Yet this example is really much simpler than the problem of air preheating, and again, the latter problem is not really made easier by concentrating on air temperature in uptakes and on the more obvious methods for making this air as hot as possible. Also to be considered are such things as the total amount of air needed theoretically for complete combustion at any selected fuel rate, together with that needed for oxygen absorbed into the bath and for burning the combustible gases given out from the bath, plus some excess air. Furthermore, how does this total air requirement relate to the regenerator efficiency at various loadings? How much of the total air equivalent in stack gases will come from preheated air passed through the whole length of the incoming regenerator and how much from air leakage into various zones of the whole furnace system? Does the effect of leakage air differ depending on whether it leaks into the ingoing regenerator, the furnace above floor level, the outgoing regenerator, or the stack zone? Over a number of years we have in this Laboratory experimented with methods for measuring hot gas temperatures; we have measured, approximately, the efficiency of a few regenerators on commercial furnaces and we have developed an electrical analogue' for calculations of regenerator efficiency. The whole problem is too complex to cover in one paper even with a more complete background of data than we possess, but we can attempt here to clarify certain aspects in a general way, hoping to assist specific plant studies on individual furnaces. Some of our earlier experience has been incorporated into chaps. 4, 19, 20, and 21 in the book "Basic Open Hearth Steelmaking,"" and on the assumption that this text is available to most readers, this discussion has been shortened by references to that book. The most recent design of a flowing-gas thermocouple which we are now using for measurement

Jan 1, 1955

By W. H., Bassett

IN January, 1926, the writer began a study of the commercial value of beryllium in its relation to copper. The purpose of the investigation was not to make a mere laboratory study of the characteristics of beryllium-copper alloys, but to determine under mill conditions the behavior of copper when alloyed with beryllium and when other metals were added to such alloys; in other words, it was considered important to know whether the alloy bars could be hot-rolled or cold-worked, and what limits in composition would have to be established for these two methods of working. Consequently, more stress has been placed on the practical application of the alloys in engineering and commercial use than on a discussion of the metallography of the specimens examined, although this part of the matter has not been entirely neglected. Credit should be first given to G. Oesterheld,1 who, in 1916, published the copper-beryllium equilibrium diagram. Fig. 1 has been copied from the original article with the simple substitution of English words. It will be noted from Fig. 1 that alpha solid solutions are formed, up to about 1.5 per cent. Be (10.5 atom.). The freezing-point curve, after falling from 4.2 per cent. Be (23.5 atom.), enters into a very complicated form over the entire beta region. There is a possible compound, Cu2Be, at about 6.5 per cent. Be (32.3 atom.). The pure gamma range extends from 11.2 to 12.2 per cent. Be. (46.5 to 49 atom.), after which the liquidus rises rapidly to 30 per cent. Be. (75 atom.). The delta solution extends from 25 per cent. to an unknown percentage of beryllium. There is a eutectoid point at 575" C., extending from 1.5 to 11.2 per cent. Be., the maximum time interval of which occurs at 6 per cent. Be. Below this line the beta solution is described as breaking up into alpha and gamma. Microscopically, the eutectiferous alloy thus formed closely resembles pearlite, and requires a high magnification for its resolution.

Jan 1, 1927

By R. W., Ludt

THE use of froth flotation for the separation of minerals has become one of the most important of ore dressing processes. Its particular adaptability to the enrichment of low grade ores has made the process an important factor in the national economy. The methods have been extended to the recovery of a great number of minerals. Among the few minerals which have resisted efforts toward industrial flotation is chrysocolla, a hydrated partly colloidal copper silicate. Chrysocolla, being a product of natural oxidation, has been found to occur in small quantities with many ores which are recovered by flotation methods. In present practice, these small quantities of copper silicate pass off with the tailings and are lost. The advantages to be gained by a satisfactory process for the recovery of chrysocolla is apparent. Any application of principles which points a way toward the satisfactory industrial flotation process for copper silicate would be of advantage. This paper presents an attack on this problem.

Jan 1, 1949

By S. Senderoff, D. E. Couch

A PROCESS for the electrolytic preparation of molybdenum from molten salts has been described previously. This previous work centered on electrowinning and electroplating characteristics of the process. In this paper the properties of the process as a refining medium are discussed. In producing metals from ores, it frequently has proven most economical to obtain a comparatively impure metal by smelting, and to purify the product by electro refining. Direct electrowinning from molten salts is rather difficult when a porous diaphragm is required in the process, as in the case of molybdenum, since satisfactory refractory materials for such a diaphragm in molten salts have not yet been developed. In electrorefining, however, a soluble anode is used. By suitable control of anode and cathode current densities, the lower-valent molybdenum compound used in the electrolyte is not oxidized, and a diaphragm is not needed, thus avoiding difficulties which are encountered with other electrowinning processes. Electrorefining may result from two different processes in the cell: Dissolution of the molybdenum and the less noble impurities with a separation of the more noble impurities as an anode sludge; the electrodeposition of molybdenum and the more noble metals at the cathode, leaving the less noble metals in the electrolyte. The amount of separation which occurs in each process is determined by the potential of the operating electrode and the relative concentration of the various species at (or in) that electrode. Since molybdenum has a comparatively noble position in the EMF series in molten alkali halides,2 most refining would be expected to occur at the cathode. In this work only the cathodic process was studied. The objective was to determine to what extent various metallic ions, if they dissolve from the anode or are otherwise introduced, were concentrated in the electrolyte. The necessary conditions for the separation of each and the limiting concentration which may be tolerated in the electrolyte for effective refining were studied. The deposits were chipped from the cathode and analyzed spectrochemically for metallic impurities. Besides determining the major contaminant, a general qualitative analysis was obtained to be sure that other metallic ions were kept at a very low level during each series of runs and to be sure that the baths were not being contaminated accidentally. Samples of the bath, for analysis, were taken at the temperature of operation so there was no chance of selective crystallization during cooling. In this way the ratio of molybdenum in the bath to the other metal could be determined readily. These data permitted a check on the anode efficiency and any volatilization that took place during the electrolysis. Experimental Apparatus—The apparatus as shown in Fig. 1 was used to determine the effect of the various foreign ions upon electrodeposited molybdenum. The entry chamber A allowed the cathodes to be removed without cooling the bath and without allowing oxygen or moisture to enter." The cathodes were made Prom 0.25-in. tungsten rods and were rotated during the electrolysis. The balloon in the lower arm B was inflated with helium after the cathode was drawn up into the upper section for removal. A new cathode was then placed into the entrance chamber A after which the chamber was flushed

Jan 1, 1959

By J. P. Nielsen

When two grains in a polycrystalline specimen meet at a point in the course of grain-boundary movements, and the new boundary created at the point is one of relatively low specific free energy, a nonequilibrium boundary condition occurs. The nonequilibrium is enhanced if the other grain boundaries involved at the point of meeting are relatively high in specific free energy. The nonequilibrium results in a correction by growth of the complex grain (two subgrains) to a large size, sufficient in many cases for it to become a self-propagating unit, i.e., a recrystallization grain. This mechanism, herein called "geometrical coalescence," is proposed as a logical origin for recrystallization nuclei, particularly for secondary recrystallization. RECENT publications1-' indicate that considerable progress has been made toward a complete understanding of the three microstructural processes in metals: recovery, recrystallization, and grain growth. However, a major problem persists, namely, the origin of recrystallization nuclei. It is generally accepted that once a recrystallization nucleus, a secondary recrystallization nucleus particularly, has reached a certain size, it will continue to grow at the expense of its neighbors, simply as a consequence of grain-boundary free-energy considerations. What has remained obscure, however, is the mechanism whereby such a nucleus comes into being and grows to the self-propagating size. Recrystallization Nucleation in a Two-Dimensional Grain-Boundary System In the left portion of Fig. 1 is shown schematically the meeting of grains 1 and 4 on disappearance of boundary 2-3, i.e., the boundary that existed between grains 2 and 3, as might happen in the course of normal grain-boundary migration. If, by chance —and the probability is not necessarily small— grains 1 and 4 are close together in their mutual orientations, or in some other special way mutually oriented so that the new common boundary has relatively a very low specific free energy (hereinafter referred to by s), then the grain boundary arrangement as depicted in the figure is highly unstable. This is particularly so if boundaries 1-3, 3-4, 4-2, and 2-1 are high in their s values. Assuming, for the sake of argument, that boundary 1-4 has a s equal to zero, and the s's of all the other boundaries present are equal to each other and constant, then the boundary arrangement on the right in Fig. 1 will be

Jan 1, 1955

CHARLES CATLETT, Staunton, Va.-A great many important things have been said in an interesting way but the most startling, to me, is that for 25 years, the price has been in the neighborhood of $1.00, with a minimum of 86 cents and a maximum of about $1.25. It is evident that during all that time the price was inadequate. The amount paid to be divided between the land owner, the capitalist, and the laborer was too limited or too irregular to permit of a healthy condition; yet it is evident that the enormous industrial development of this country was fostered by this very cheap coal, probably the cheapest that has been provided in quantity in any country. I entered the coal business when the price was lowest and my interest and sympathy went out to the man who spent his days in the dirt and foul air of a mine and I was filled with schemes for helping. But at the end of the year, after exercising every economy and in blissful ignorance of proper depreciation and obsolescence, there was still only a profit of 1 cent per long ton for what I think is one of the best coals in the world, the best grade of New River coal. We are now paying a great deal more than we did a few years ago and probably will continue to pay more for some time; but any one who thinks we will not have low prices again is very young or very forgetful. Yet we must have larger profits if the industry is to be healthy, if the three parties to the production are to get adequate returns and be in shape to provide the improvements which will be needed to take the place of and supplement the man-power. This will come when the men are making so much money that the operator wants to replace the man by a machine, and when the operator is making so much money as to be in a shape to make the improvements. He will not put in machinery unless the machinery will save money. We have got to pay a higher prime cost for coal, but engineering skill will teach us how to handle it better and use it more economically, and the ultimate result ought to be of advantage to the public.

Jan 3, 1920

By George Kunz

(Butte Meeting, August, 1913.) . THE manifold inconveniences resulting from the absence of a uniform standard of mass for determining the weight of precious stones have long been obvious. This lack has been keenly felt in commercial transactions, and those who have devoted time and research to the study of historic diamonds and precious stones have had frequent occasion to deplore the absence of such a standard in the past. In a paper read in Chicago in 1893, before the International Congress of Weights and Measures, held in connection with the World's Columbian Exposition, the writer suggested dividing the carat into 100 parts, and constituting a standard international carat of 200 mg. ; that is, 5 diamond carats or 20 pearl grains to a French gram, making 5,000 carats or 20,000 pearl grains to a kilogram. He also called attention to the fact. that while this would depreciate the present diamond carat or pearl grain only about 2.5 per cent., it-would abolish the troublesome discrepancies between the various carat-weights now in use, and could be easily explained and understood everywhere.1 The subject of the various diamond carats, their incongruity, and the resulting confusion as to the correct. weights of historic gems when definite records of them are searched for, has been fully treated by me in an extensive study of this subject in The Book of the Pearl.2 To the earnest and unremitting efforts of C. E. Guillaume, Director of the Bureau Internationale des Poids et Mesures at Sèvres, is largely due the eventual success of this eminently desirable reform, which he has constantly urged both by articles on the subject and by addresses delivered in Paris before the International Committee for Weights and Measures.2

Jan 7, 1913

By W. J. Putnam, H. F. Moore

THIS paper gives a preliminary summary of results of tests on the resistance to fatigue under reversed stresses of steel subjected to cold-working and of tests to determine the effect of rest on the endurance of steel under reversed stresses. The tests of cold-worked steel were undertaken as part of an investigation by the National Research Council Committee on Fatigue Phenomena of Metals and were made by Mr. Moore; the tests to determine the effect of rest are part of an investigation now being made by Mr. Putnam. Both series of tests were made in the Materials Testing Laboratory of the University of Illinois. The preliminary results of these tests are reported now because it is believed that test data on fatigue strength should be available at the earliest possible moment; however, it should be clearly kept in mind that this paper gives preliminary results and is intended to encourage discussion rather than to give final conclusions. Before giving the data of the tests, the writers will explain the term "fatigue strength" as used in this paper and will discuss briefly the method used in interpreting the test data of repeated-stress tests. " The term fatigue strength, as applied to a material, may be defined as that maximum fiber stress that will cause fracture of the material after the repetition of any given number of cycles of stress, each cycle involving the variation of stress from a minimum to the above maximum. The fatigue strength, then, varies with the number, of repetitions occurr,ing in the lifetime of a structural or machine member and with the range of stress for each cycle of stress. The fatigue strength will be lower for a member in which the stress is completely reversed than for one in which the stress varies from zero to a maximum. In interpreting test data from fatigue tests under reversed stress, the stresses and corresponding numbers of repetitions required for failure are plotted on logarithmic cross-section paper. In 1910, Professor

Jan 1, 1920

By Nathaniel Arbiter, Norman L. Weiss

Factors now accelerating the trend to larger concentrators and larger equipment units are reviewed. After almost 40 years of stability with unit sizes less than 100 cu ft, 200 and 300-cu-ft flotation-cell units are now operating and a 500-cu-ft machine is being built. The design of these cells is considered, as is the design and operation of flotation circuits against the background of the need to achieve greater process efficiency while holding down operating and investment costs. Particular attention is paid to the scale-up problem. With the newer cells having five to ten times higher unit capacities, the advantages resulting from this in sampling, mill operation nnd control, testing and plant layout are indicated, as well as the broad direction for changes in the overall design of 50,000 to 100,000-tpd concentrators. A comparison is given of the available larger cells of U.S. manufacture with analysis of differences and similarities. After almost 60 years of application in this country, froth flotation is still by far the most important process for concentrating metallic ores as well as a number of nonmetallic ores including fluorspar, phosphate rock, and potash. Its use on iron ores and coal is increasing. In the metals groups, the largest plants and the greatest progress are found in copper and molybdenum mills, and the section of this paper that deals with applications will be found most pertinent to that field. During its long reign, the flotation process has passed through several phases of evolution and has achieved progress in economy, efficiency. and simplicity of operation. Much of today's success in the flotation of low-grade ores can be attributed to that progress. Even more important to us than our ability to cope with today's conditions are tomorrow's problems, which we can contemplate with some prescience and confidence. Our annual production of copper, lead, zinc, and molybdenum—to mention only the most important metals occurring in the U.S. as sulfides- must be held at a high level in the face of a declining grade of ore. It is probable that the tonnage of such ores mined and processed will double in this country before the turn of the century. Not only will many of the existing flotation mills be enlarged again and again, but larger mills will be built here and abroad to meet our needs. At this moment, in fact, there are at least three proposed new foreign mills or expansions that envision 100,000-tpd milling rate on copper ores. Table 1' shows the trend in copper ore grades in U.S. in recent years. Table 2' shows the projected annual growth rate of consumption 1966-1985. A study by one of the authors 25 years ago showed that of 32 copper mills in North America with more than 100-tpd capacity, only eight had capacities exceeding 10,000 tpd. The average for these eight was 25,000 tpd. These included the three giants of that day, Magna, Arthur, and Morenci; the other five averaged a little over 10,000 tpd. Today instead of the eight mills with capacities exceeding 10,000-tpd capacity we have 25, and the mill with less than 1000-tpd capacity is fast becoming extinct. Much less range is found in lead and zinc mills, which are still generally small except for the few with over 1000-tpd capacity in Missouri, Utah, and Idaho, and a few zinc ore mills in the same size category. Molybdenum and potash operations are the fastest growers in North America today, but when one considers the projected increase in consumption of copper together with the projected decrease in grade of ore (3% annually), he can foresee that copper ore milling will soon lead the list. Technological improvements to meet the challenge of lower-grade ores have taken the wholly predictable path toward bigger equipment. In mining, shovels, draglines, scrapers, and trucks have led the way; in ore processing, we have seen a rapid increase in size of grinding mills, flotation cells, and pumps. Twenty-five years ago, for example, the 10-ft-diam ball mills treating 2000 tpd at Morenci were considered the giants, but today some grinding mills in similar service have five times that capacity. The specific matters to be dealt with in this paper are 1) the underlying principles that affect flotation-machine design, and the conditions required to produce

Jan 1, 1971

By Donald W. Gentry

The recently established SME-AIME Education Board has completed its first full year of operation, and its six constituent committees-Accreditation, Career Guidance, Continuing Education, Educational Statistics, and Scholarship-have been extremely active in identifying and initiating programs to serve the educational needs of the Society of Mining Engineers and its members. The Continuing Education Committee-primarily concerned with the needs and problems associated with providing a continuing education program for SME members-is responsible for: (1) determining the need for short course offerings; (2) developing needed courses, along with new means and opportunities of presentation; and (3) scheduling and evaluating course presentation. In response to this charge, the committee has sponsored two short courses at the SME-AIME Fall Meeting in St. Louis and three short courses at the AIME Annual Meeting in Denver. Its major effort, however, has been directed toward the 1978 SME-AIME Fall Meeting to be held in Orlando, Fla., where 18 days of short-course instruction have been scheduled. A portion of these courses will be presented prior to the meeting, with the remainder to be presented in Nassau, Bahamas, immediately following the meeting.

Jan 7, 1978

By Clinton H. Crane

IN October, 1925, J. R. Finlay delivered an address entitled, "The Future Price of Lead." Lead was then selling at 8.85c. and Mr. Finlay and most of the rest of us were concerned about the shortage. Now, a little more than six years after, lead is selling for 3.55c., and nobody is worrying about the difficulty of getting plenty of it. For 1931 the U. S. G. S. reports United States produc¬tion as 390,000 tons of recoverable lead in ore. Our estimate is slightly higher than this; nearer 400,000 tons. This is the smallest production since 1911 and 1912, but in spite of this, stocks for the year-have increased nearly 50,000 tons. Our estimate for the production outside the United States is 1,120,000 tons, or more than double the production of either 1920 or 1921, and 30 per cent more than in 1913. Stocks abroad probably have not increased more than stocks in the United States. In other words, foreign consumption is double what it was in 1.920-21, whereas domestic consumption is substantially less. It looks as if Mr. Krueger, president of the Swedish match trust, was right when he said that "This country had better stop worrying about Europe and think a little bit more about its own affairs."

Jan 1, 1932

CONTENTS PAGE WARNER, ROBERT K.-Efficiency of Screening. Discussed by Robert C. Canby,Robert K. Warner, W. 0. Borcherdt, V. E. Flanagan 1 FAHRENWALD, A. W.-Surface Reactions in Flotation. Discussed by Arthur F. Taggart, A. M. Gaudin 7 Efficiency of Screening Discussion of the paper of ROBERT K. WARNER, presented at the New York Meeting, February, 1924, and issued, as Paper No. 1293-M, with MINING AND METALLURGY, January, 1924. ROBERT C. CANBY, Wallingford, Conn.-On page 2, the author speaks of the importance of using a testing sieve clothed with the same material as the screen being studied; does he regard this as essential? In practice, how does the screen efficiency calculated by the difficult-grain method compare with those computed by the usual undersize method? Is it possible by the author's method, to obtain a measure of efficiency that can safely be compared with that of another screen sizing a different ore? ROBERT K. WARNER.-The use of a testing sieve clothed like the mill screen being studied is desirable but not essential for routine work about a mill where the character of the feed varies little, as regards the shape of the grains and the sizing test. Close agreement in size of aperture between the sieve and the screen is, however, important if the data obtained are to be valuable to others working on a different feed. If the testing sieve is coarser than the mill screen, the efficiency will appear unfairly low; whereas if it is too fine one might get efficiencies greater than 100 per cent. because of the apparent oversize in the undersize. In either of these cases the usefulness of the difficult-grain method of calculation would be reduced. As screens wear, the undersize grows appreciably coarser so that it is desirable to test screens only when the cloth is new or else to provide testing sieves that are worn about the same amount. The futility of making efficiency observations on a. distorted or broken screen surface is obvious.

Jan 6, 1924

By D. L. Ritter, N. J. Grant, B. C. Giessen

Forty-six alloys covering the complete concentration range of the Nb-Rh system were examined by metallographic and X-ray methods; solubility limits of terminal ad intermediate phases and transformation and solidus temperatures were determined. There are nine intermediate phases; two eutectic, five peritectic, two eutectoid, and three peritectnid reactions were encountered. The experimental methods and the constitution diagram are described in Part I. Part 11 deals with the crys -tallography and relation of the intermediate phases. The Nb-Rh system was selected for study for three reasons: first, it was hoped that the anticipated similarity of the Nb-Rh phase diagram to the Ta-Rh1 and Ta-Ir2 diagrams treated earlier (see also Ref. 3) and to the Nb-1r4 diagram treated simultaneously would allow a crystallographic and comparative evaluation of the intermediate phases in these related systems; second, it was likely that the previous experimental methods could be carried over without significant change; and third, it was expected that the interesting mechanical properties of the a1 phase of the Ta-Rh and Ta-Ir systems would also be encountered in similar phases of the new systems, and that these phases could be investigated more closely. The phase diagram had not been treated fully in the literature; only brief surveys of some intermediate phases were found.5-7 It was attempted to cover the details of the phase boundaries, Part I, as well as to present the crystallography of the intermediate phases, Part 11. EXPERIMENTAL METHODS Specimen Preparation and Treatment. The following high-purity starting materials were used. 1) Niobium powder, 99.8 pct pure, mesh size -60 +200, from Wah Chang Corp., Albany, Ore. The analysis furnished by the supplier showed the following major impurities: Ta, <500 ppm; Zr, <500 ppm; 0, 300 ppm; Fe, 280 ppm; Ti, <I50 ppm; W, Si, <100 ppm; C, N, 30 ppm; all other impurities, <30 ppm. 2) Rhodium powder, ground sponge, 99.9+ pct pure, mesh size -80, from J. Bishop Co., Malvern,

Jan 1, 1964

By F. E. Bash

About three years ago, the writer presented a paper1 on the rate of heating and cooling of a 24-in. round ingot. The present paper deals with work done on larger ingots at the plant of the Allis Chalmers Mfg. Co., Milwaukee, Wis., in August, 1919. Sufficient data on ingots of all sizes were desired to enable one to calculate in advance the rate at which any ingot could be raised to a forging temperature without injury to the steel, determining also the temperature gradient between different sections of the ingot and the length of time necessary to heat to the center. This has now been experimentally determined for 24-in. and 45-in. ignots by the writer and for a 18-in. cube by E. F. Law.2 J. F. Harper has also made tests on heat-treating temperatures on 30-in, ingots; which data are included in this paper. No attempt is made here to work out a formula for the heat penetration of steel, for too many factors are involved concerning which we have not sufficient information, as for instance, the specific heat of steel up to 2300&apos; F. It is hoped that eventually a set of formulas will be developed so that one may calculate maximum allowable temperature differences within an ingot for any particular kind of steel and the rate of absorption of heat and temperature gradient within the ingot for a definite temperature head. EXPEEIMENTAL INGOT The ingot tested was 45 in. in diameter, having been forged into the shape shown in Fig. 1 from 54 in. diameter. Seven -in. holes were

Jan 1, 1923

By F. E. Bash

About three years ago, the writer presented a paper1 on the rate of heating and cooling of a 24-in. round ingot. The present paper deals with work done on larger ingots at the plant of the Allis Chalmers Mfg. Co., Milwaukee, Wis., in August, 1919. Sufficient data on ingots of all sizes were desired to enable one to calculate in advance the rate at which any ingot could be raised to a forging temperature without injury to the steel, determining also the temperature gradient between different sections of the ingot and the length of time necessary to heat to the center. This has now been experimentally determined for 24-in. and 45-in. ignots by the writer and for a 18-in. cube by E. F. Law.2 J. F. Harper has also made tests on heat-treating temperatures on 30-in, ingots; which data are included in this paper. No attempt is made here to work out a formula for the heat penetration of steel, for too many factors are involved concerning which we have not sufficient information, as for instance, the specific heat of steel up to 2300&apos; F. It is hoped that eventually a set of formulas will be developed so that one may calculate maximum allowable temperature differences within an ingot for any particular kind of steel and the rate of absorption of heat and temperature gradient within the ingot for a definite temperature head. EXPEEIMENTAL INGOT The ingot tested was 45 in. in diameter, having been forged into the shape shown in Fig. 1 from 54 in. diameter. Seven -in. holes were

Jan 1, 1923

By D. J. Weintritt, J. C. Cowan

Methods of studying oilfield mineral .scale deposition in the laboratory do not work for barium sulfate because only small nonadhering crystals ore formed. On the other hand, barium sulfate scale found in down-hole or surface equipment is strongly adhering and may contain very large crystals. Results suggest that most of the difference derives from the extremely low .solubility of barium sulfate. Firm adherence of scale and the consistent development of oriented crystals 100 microns and larger suggest a relationship between scale adherence and crysta1 growth. Data from this study indicate some reasons for barium sulfate&apos;s occurring as a deposit in oilfield waters. The unique characteristics as well as the associative properties of barium sulfate scale as related to calcium carbonate and calcium sulfate ere shown. INTRODUCTION The first observed deposition of barium sulfate scale in oilfield production equipment is unknown but probably coincides closely with the beginnings of the oil industry itself. Moore&apos; described barite oolites in some producing wells in the Saratoga field, Tex., in 1914. Currently, positive identification of barium sulfate deposition is recorded in most of the major oil-producing areas of the United States. As efforts continue to identify the composition and functions of the materials found in scale deposits, the geographic and economic importance of the part played by barium sulfate will become even more important. This belief is based on the following Facts. 1. Many oil fields in the U.S. are in or entering their mature phase and the volume of produced water is increasing; thus, the cumulative effects of low-solubility minerals become more significant. 2. Secondary recovery techniques and increased emphasis on waste disposal systems require controlling the behavior of all constituents of water to prevent formation plugging. 3. The more general use of instrumental analysis has reduced the time and improved the accuracy of complete scale and water analysis. 4. Once formed, barium sulfate scale is resistant to present methods of chemical removal; therefore, costly mechanical methods are necessary. Thus, new methods of prevention and control will require specialized chemical techniques and better knowledge of the causes of scaling. Consider the results of some simulated field deposition studies illustrating this need (Fig. 1). The studies utilized a Scale Deposition Test Cell* and each coupon was exposed to 8,000 ml of a 300 me/liter solution of the various types of mineral scale. All tests were conducted at 140F. Only the calcium sulfate and calcium carbonate deposited scale. The solution containing 300 me/liter of barium sulfate exhibited copious amounts of precipitate, but none of the precipitate adhered. The fluids from the effluents were also allowed to impinge on glass slides. Fig. 1 shows the relative adherence of the three types of scale on unetched glass. The barium sulfate failed to deposit in the same manner as the calcium carbonate and calcium sulfate solutions.

By G. V. Woody, J. D. Price

Many producers of by-product coke have spent considerable time and given considerable thought to the use of a substitute for low-volatile coal as an admixture with high-volatile coking coal for charging high-temperature by-product coke ovens. Generally speaking, there are two principal reasons for this, one being occasioned by the lack, in certain localities, of low-volatile coal delivered at a cost sufficiently low to permit its use, and another being the desirability of finding a substitute product with a delivered cost lower than the cost of the low-volatile coal. It was for the first of these reasons that the Colorado Fuel and Iron Corporation was particularly interested in this subject. This paper deals with the experimental work and the results obtained by that company in producing and using low-temperature char. Overcoming Defects in Coke from Colorado Coals Colorado coking coals are all of the high-volatile type and were all laid down during the Cretaceous period, which makes them something like 80 million or so years younger, geologically, than eastern coals. They behave, when coked alone, something like the Pittsburgh-seam coals. They make a very brittle, highly cross-fractured, fingery coke of fair shatter value but quite low in resistance to abrasion. Mixed with a low-volatile coking coal such as Poca- hontas or certain Oklahoma coals, they make a coke that is equal to virtually any of the eastern cokes. The inferior quality of Colorado coke has long been recognized, and a considerable amount of time, effort and money has been spent upon investigations of means for improving its physical properties. While it has been found that some improvement in quality of coke can be secured by the adjustment and control of such variables as pulverization, moisture content, oven temperature and bulk density of charge, and through the addition of certain inert materials such as pulverized coke breeze, pitch, high-volatile noncoking or semicoking coal, the benefits so gained have been comparatively small. There are no low or medium-volatile coking coals available within a reasonable freight-cost distance. Therefore a substitute for low-vol-atile coal was developed through low-tem-perature carbonization—or, more correctly speaking, through the partial devolatiliza-tion—of high-volatile coals. The blending of this lower volatile char with the high-volatile coking coal distinctly improves the quality of coke made from Colorado coal. There is nothing especially new about this idea. In 1908 a Japanese patent was taken out by Kotaro Shimomura, whose claim was: The method of making a nonfingery coke out of bituminous coals, without using natural coals of low-volatile matter; according to which a certain coal is heated at a temperature between about 300°C. and about 600°C. so as to leave about 15 to 25 per cent volatile matter in the coal; this coal is mixed with the

Jan 1, 1944