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By John J. Schanz

INTRODUCTION The two prior Seeley W. Mudd books on mineral economics devoted their final chapters to themes having to do with the future. In 1932, the last chapter in Mineral Economics8 was "The Future Value of Mineral Property" by J. R. Finlay. Both the first edition (1959) of Economics of the Mineral Industries7 and the slightly revised second edition (1964) closed with "Minerals and Man's Future" by Zay Jeffries. We are not inclined in this third AIME volume on mineral economics to ignore the tradition of the prior volumes. Moreover, it always is fitting and proper to close a book about the mineral industries by examining in a rather general way the challenges and uncertainties that are perpetually strewn in our path as we pick our way into the future. Unlike its predecessor volumes, which were written to acquaint the nonspecialist with the economic characteristics of mineral resources and to provide data about the mineral industries, this edition was also designed with the mineral economics student and practicing specialist in mind. Thus it is anticipated that it will be both a reference as well as suitable for use as a textbook. Consequently, this closing examination of our changing perspectives about mineral resources devotes some space to the past events and current status of mineral economics as a field of study. The fears and forecasts of the past can be a perceptive and usually deflating way to begin an examination of the future. As we look up and down the track of time from the 1975 platform, we find that many of the past views of the future were transient in character, others were imperfect in their prognosis, while others were based on fears that now seemed unfounded. But down through the years some of the observations or expressions of concern have a persistency which perhaps reflects some elemental characteristics of mineral resources. On these "truths" rest our contemporary understanding of the economics of the mineral industries. The uncertainties and priorities for the next 25 or more years, not surprisingly, have a somewhat different mix than those that seemed most compelling to our predecessors in 1932 or 1959. So it is appropriate that we review how the perspective of the mineral economist has changed over the span of a half century. In light of the particular interest of AIME on the role technology must play, it is fitting that we close with some observations on that topic. THE EMERGENCE OF MINERAL ECONOMICS AS A FIELD OF STUDY The publication of Mineral Economics by the AIME in 1932 as a Seeley W. Mudd volume marked the end of emerging interest in

Jan 1, 1976

By L. H. Levenson, Charles S. Barrett

Numerous determinations1-' of the texture of cold-rolled polycrystalline iron, steel, and ferritic alloys have been made with good agreement among the various observers as to the principal features of the texture common to all, but there has been no adequate explanation of the texture in terms of, the behavior of individual grains. This study presents an explanation of the texture based on the observed rotation of single crystals when deformed in a manner similar to a grain in the aggregate, a method we have found useful in leading to a better understanding of compression and tension textures;8,9,10 in particular, the aim has been to determine experimentally the end orientations of rolled single crystals of iron. These are discussed in connection with the point of view of Weverl and of Boas and Schmid2 that the rolling texture is a simple compression texture (compression normal to the rolling plane) superimposed on a simple tension texture (tension in the rolling direction), and the point of view of Kurdjumow and Sachs3 that three ideal end orientations exist. Experimental Method The single crystals used were in the form of disks cut from a stock of crystalst pre- pared in this laboratory by a strain-anneal method from mild steel decarburized and purified in hydrogen, the details of which have been described elsewhere.11 To ensure maximum elongation in the rolling direction and to simulate approximately the rolling of a grain in an aggregate, lateral flow was constrained by placing each crystal in a close-fitting hole in a plate of copper or iron of equal thickness and rolling the plate and crystal together. About one-third of the crystals were rolled in a "sandwich" pack by adding top and bottom cover plates. The two-high mill used had 8-in. diameter rolls and was operated at 15 r.p.m. The initial diameter of the crystals was between 1/4 and 1/2 in., while the starting thickness varied from 0.031 to 0.376 in. With a single exception, which was rolled 63 per cent, the crystals were rolled to a final reduction in thickness of 82 to 97 per cent in 20 to 40 passes, alternately changing the sense of the rolling direction. The initial orientations of the crystals were determined either by X-ray back reflection12 or by optical reflectionl3 from etch pits; the mean final orientations were determined by X-ray Laue photograms using molybdenum, silver, or tungsten radiations, by optical reflection, and in many cases by both methods. Results Twenty-six crystals were studied. It was found that there were three typical be-

Jan 1, 1941

By Harlowe Hardinge

[Crushing Tests by Pressure and Impact (T.P. 1895, by F. C. BOND,, Min. Tech. Jan. 1946). Discussion by H. HARDINGE and the author 1 Pumping Sands, Slurries and Slimes and Installation and Performance of Sand Pumps (T.P. 2013, by W. B. STEPHENSON, Min. Tech., May 1946 and T.P. 1978, by C. G. SOUTHMAYD, Min. Tech., March 1946). Discussion by OSCAR LEE, G. J. HOLT, L. D. ANDERSON, and the author 2 Rapid Polarographic Method for Determination of Copper, Zinc and Lead in Flotation Tailings (T.P. 2010, by G. A. CROWE and W. T. BISHOP, Min. Tech., March 1946). Discussion by E. H. ROSE and the authors 4 Ball Wear and Functioning of the Ball Load in a Fine-grinding Ball Mill (T.P. 1984, by W. I. GARMS and J. L. STEVENS, Min. Tech., March 1946). Discussion by J. F. MYERS, E. W. DAVIS, F. M. JARDINE, H. HARDINGE, F. C. BOND and the authors 5 Correlation between Mineral Behavior in Cataphoresis and Flotation (T.P. 2005, by A. M. GAUDIN and S. C. SuN, Min. Tech., May 1946). Discussion by 0. C. RALSTON, G. L. SIMARD and D. J. SALLEY 7 Accident Experience in Milling (T.P. 1981, by J. F. MYERS and S. E. SHARP, Min. Tech., May 1946). Discussion by C. E. MCKNIGHT, C. H. BENEDICT, T. S. WILSON and S. POWER WARREN9 Mill Design for Labor Economy (T.P. 1997, by NORMAN WEISS, Min. Tech., May 1946). Discussion by R. S. HANDY and the author13] Crushing Tests by Pressure and Impact (0.C. Ralston presiding) HARLOWE HARDINGE.*-Has the impact method described in the paper been used to compare the resistance to reduction in size of various materials with actual operation results of mills in the field? Is this impact method a better index of the true resistance to grinding under actual field conditions than other methods the author has reported in the TRANSACTIONS? F. C. BOND (author's reply).-The impact crushing device was designed to measure relative resistance to crushing, or what may be called the crushability, and is not used for comparing the resistance to grinding, or grindability. The correlation between crushability and grindability of different materials is not at all close, since fractures, zones of weakness, and structural features of rock have a much greater effect upon the crushability than upon the grindability. We still depend upon our standard grindability tests for measuring resistance to grinding, and use the impact crushing tests only in relation to crusher installations. The impact crushing device has several advantages over the measurement of crushing strength in pounds per square inch as an index of crushing resistance.

Jan 1, 1946

Discussion of the papers of FREDERICH LAIST and HAROLD W. ALDRICH, FREDERICH LAIST and F. F. FRICK, W. L. AUSTIN, and STUART CROASDALE, presented at the Salt Lake meeting, August, 1914, and printed in Bulletins No. 91 and 92, July, and August, 1914. R. C. CANBY, Wallingford, Conn. (communication to the Secretary*). -Apropos of the experimental reduction of copper from cuprous chloride by fusion with ground limestone and coke, as described by Messrs. Laist and Frick, and also proposed for use at Chuquicamata., it might be of interest to refer briefly to the experience at Argentine, Kan., in the smelting of the suboxide of copper (as produced in the Hunt and Douglas process) containing appreciable quantities of chlorides. In the Engineering and Mining Journal of June 10, 1911, I. gave some account of what had appeared to be crucial difficulties with the Hunt and Douglas process, and how the filter-press difficulty had been so rapidly overcome after my discovery of the effect of the sodium salts in the leaching solutions. I also described the effect of the cupola furnace gases upon the lime rock with which I had replaced the coke in the absorption tower, which apparently showed the presence of considerable-chlorine, clue to the imperfect conversion of the subchloride to suboxide, as well as that clue to the difficulties of washing thoroughly the suboxide with the means provided, and also to the presence of chlorides in the cement copper which was smelted With the suboxide. The fumes from the cupola, when issuing direct into the atmosphere, were very dense and white, except when a hot furnace top caused a reddish tinge due to dust. These, fumes had a peculiarly suffocating effect, so that when they were blown over into the town the population suffered the greatest actual distress, not the mere annoyance which would have been clue to a disagreeable odor. The absorption tower had been built at the instance of the city authorities to abate this nuisance, and not with the idea of making any additional metal savings. A serious feature in connection with the Argentine experience with the Hunt and Douglas process was, that it had been implicitly assumed that there could be no metal loss in the Hunt and Douglas process itself, and that all attempts to locate the ruinous copper losses which were accumulating in the metallurgical accounts were devoted to the roasting department. It is possibly unfortunate for metallurgical advance along these particular lines that just about the time of this discovery of this previously unsuspected volatilization loss, in the smelting of the suboxide, the Kansas City company made a contract with the Western Union Telegraph Co. for its supply of copper sulphate and the Hunt and Douglas

Jan 11, 1914

By Charles S. Barrett, L. H. Levenson

Numerous determinations1-' of the texture of cold-rolled polycrystalline iron, steel, and ferritic alloys have been made with good agreement among the various observers as to the principal features of the texture common to all, but there has been no adequate explanation of the texture in terms of, the behavior of individual grains. This study presents an explanation of the texture based on the observed rotation of single crystals when deformed in a manner similar to a grain in the aggregate, a method we have found useful in leading to a better understanding of compression and tension textures;8,9,10 in particular, the aim has been to determine experimentally the end orientations of rolled single crystals of iron. These are discussed in connection with the point of view of Weverl and of Boas and Schmid2 that the rolling texture is a simple compression texture (compression normal to the rolling plane) superimposed on a simple tension texture (tension in the rolling direction), and the point of view of Kurdjumow and Sachs3 that three ideal end orientations exist. Experimental Method The single crystals used were in the form of disks cut from a stock of crystalst pre- pared in this laboratory by a strain-anneal method from mild steel decarburized and purified in hydrogen, the details of which have been described elsewhere.11 To ensure maximum elongation in the rolling direction and to simulate approximately the rolling of a grain in an aggregate, lateral flow was constrained by placing each crystal in a close-fitting hole in a plate of copper or iron of equal thickness and rolling the plate and crystal together. About one-third of the crystals were rolled in a "sandwich" pack by adding top and bottom cover plates. The two-high mill used had 8-in. diameter rolls and was operated at 15 r.p.m. The initial diameter of the crystals was between 1/4 and 1/2 in., while the starting thickness varied from 0.031 to 0.376 in. With a single exception, which was rolled 63 per cent, the crystals were rolled to a final reduction in thickness of 82 to 97 per cent in 20 to 40 passes, alternately changing the sense of the rolling direction. The initial orientations of the crystals were determined either by X-ray back reflection12 or by optical reflectionl3 from etch pits; the mean final orientations were determined by X-ray Laue photograms using molybdenum, silver, or tungsten radiations, by optical reflection, and in many cases by both methods. Results Twenty-six crystals were studied. It was found that there were three typical be-

Jan 1, 1941

By R. S. Moehlman

IN recent years, exploration and development by diamond drilling has become a popular attack on the problem of finding new ore shoots and developing extensions of known shoots. Diamond drilling has been used in recent years in extensive sampling of disseminated copper deposits, in preference to churn drilling. At the Harvard gold mine, near Jamestown, California, Philip R. Bradley, Jr., general superintendent, reports that: "The Harvard work was a singular example of the reliance that can be placed on diamond drilling in evaluating an ore body." Drilling at the Harvard mine will be discussed in a subsequent section of the paper. So popular has diamond drilling become, indeed, that many who have had slight experience with drilling are inclined to regard it as a cheap cure-all. This paper proposes to show some of the limitations on the application of drilling and some of the steps that should be followed in drilling programs to obtain reliable results. COST OF DRILLING Drilling has the advantage, over shaft sinking and drifting, of lower cost and greater speed. As compared with churn drilling, the advantage of bringing up a coherent sample instead of fine cuttings for a sample is obvious. Diamond drilling is generally more expensive, however, and more engineering effort is required in combining the average of core and sludge samples. Exploration of drift walls by hammer drills to a depth of less than 50 ft. is generally cheaper than diamond drilling, but beyond this depth the diamond drill is cheaper and more efficient. The principal factor in determining cost of drilling is the character of the ground drilled. Broken ground with many fissures is most expensive to drill. Hard ground is less of a problem with diamond drills than with ordinary machine drills. Hardness or softness affects only the speed of drilling. Caving, easily washed, sandy or clayey ground requires continual cementing and casing, which is the greatest cost problem. Thus, soft, incompetent ground is more costly to drill than hard, massive ground. Cost of drilling per foot in good to fair ground varies from less than $1 per foot, achieved in some districts in Ontario, the Mother Lode, and Missouri, to $5 per foot in many western states. In bad drilling ground, costs may range from $5 to $15 or more per foot, thus approaching the cost of drifting. Other factors that increase the cost per foot are: I. Short total length of hole drilled, so that the cost of moving in crews and equipment must be spread over a relatively few feet. 2. Cost of drilling per foot increases with depth beyond about 500 ft., as more time is consumed in pulling rods. 3. Cost of preparing drilling stations, getting air underground to stations, etc. 4. Inaccessibility and lack of water on the job make for higher costs.

Jan 1, 1946

The Hadsel Mill. By R. G HALL (Trans, vol 112 4000 words) This mill is designed to effect the entire reduction of ores by combined crushing and grinding from mine run size to material sufficiently fine for the usual operation of flotation or cyaniding The work is begun and completed in the same machine. The machine is designed on the principal of the elevator wheel, and crushing and grinding are accomplished by dropping a mixture of ore and water upon specially placed "breaker plates" in such a manner that the resultant material passes back into the buckets of the wheel for further elevation One method of removal of the finished product is to permit an overflow of wet pulp on one side of the mill through coarse screens A recent improvement provides for the removal of a pulp by a launder part way up the path of the buckets In either case the pulp passes to a classifier, and the oversize back to the mill in the manner familiar in ball-mill practice Although hardly through the development stage, the machine has given a very good account of itself in simplification of plant, by elimination of coarse and fine crushers with consequent reduction of installation cost and power consumption As compared with the ball mill alone, there is a reduction of power consumption as well as a material reduction of grinding steel There is an increase in cost of maintenance other than grinding steel, but as the machine passes out of the development stage there is a show¬ing of reduction of costs on this item It seems probable that with further develop¬ment and strengthening of the weaker parts further improvement may be made Ball Milling. By A. M Gow, M GUGGENHEIM, A. B. CAMPBELL AND W. H COGHILL. (Tech Pub 517; also Trans, vol 112. 24,000 words) This paper outlines a new philosophy in ball milling by using power as an index of mill operation and ascertaining the effects of the set and induced variables By recognizing the amount of ore in the mill as an important induced variable and recognizing that the amount of ore in the mill depends principally on the size of the discharge trunnion, ball volume, rate of feed, and speed of mill, obstacles formerly met were surmounted When the amount of ore in the mill is dealt with intelligently and the applied work is proportional to the amount of feed, series testing, which would otherwise give a hotchpotch, gives concordant results The study of ball paths and trajectories was seen to be fruitless when it was shown that the old attrition theory for slow mill speed

Jan 1, 1935

By Walter Crafts, C. M. Offenhauer

IN the course of study of the heat-treatment of low-alloy steels, the behavior of alloy carbides at subcritical temperatures was found to vary from that indicated by published investigations. In order to establish the ranges of formation of carbides in low-chromium steels, a series of steels containing up to 0.60 per cent carbon and 7.5 per cent chromium has been investigated with respect to the type of carbide formed after quenching and tempering at subcritical temperatures. The types of carbide formed as products of austenite transformation to pearlitic and pseudomartensitic structures have also been determined. There have been numerous investigations of the carbides in chromium steels and diagrams of the iron-chromium-carbon system.1 Westgren, Phragmén and Negresco2 identified the following carbides in chromium steels: [Cementite Fe3C (containing up to about 15 per cent Cr) Trigonal carbide Cr7C3 (containing up to about 55 per cent Fe) Cubic carbide Cr4C (or Cr23C3) (containing up to about 25 per cent Fe) Orthorhombic carbide . Cr3C2 (containing only a few per cent Fe)] The most complete information on the iron-chromium-carbon system was supplied by Tofaute, Sponheuer and Bennek3 and by Tofaute, Küttner and Büttinghaus.4 By means of dilatometric and microscopic studies, sections of the diagram were established. This diagram showed the presence of the special carbide Cr7Ca in steels containing about 2.5 per cent to 12 per cent chromium in low-carbon steels and from about 5 to 20 per cent chromium in r per cent carbon alloys from A1 to room temperature. The present investigation has shown that a carbide inversion from Fe3C to Cr7Ca occurs in chromium steels containing more than about 1 per cent chromium after hardening and tempering at subcritical temperatures above about 500°C. The type of carbide observed to be more stable in quenched and tempered specimens was also found to be produced by direct austenite transformation in corresponding ranges of temperature and composition. EXPERIMENTAL PROCEDURE The steels used in these experiments were made in magnesia-lined crucibles in high-frequency furnaces and were poured into 2-in. ingots. The ingots were forged into 1-in. round bars, and all heat-treatments were conducted on specimens of this section .size. Sections of the bar stock 8 in. long were hardened in air, oil or water depending on the composition, after heating to a temperature sufficiently high to dissolve the carbides. The low-carbon and low-chromium steels were not entirely martensitic after quenching, and contained some ferrite and pseudomartensite. No pearlite was observed in the microstructures after auenchine. The quenched bars were cut into

Jan 1, 1942

By F. R. Archibald, C. M. Nicholson

The present paper may be considered complementary to an earlier contribution on the same subject by F. R. Archibald and C. F. Jackson.1 It is particularly concerned with engineering and technological aspects of the operations at the Harleyville Alumina Plant during its period of operation from January 1945 to July 1946. The plant was built for and operated by Ancor Corporation, at Harleyville, South Carolina, as part of the government sponsored wartime program related to alumina production from domestic raw materials. The main purpose of the program was to secure technical and engineering data, based on actual plant experience, that might serve as a foundation for large scale production of alumina from raw materials other than bauxite. The Harleyville plant was directed specifically to treatment of clays found in very extensive deposits in the Carolinas and Georgia using the lime-sinter process developed by Ancor Corporation. It is considered that the purpose stated above has been served and this paper is to record and discuss some of the more important findings. The main factor in location of the plant at Harleyville was the presence in that locality of a rather remarkable deposit of high calcium limestone. The clay mining operation was centered near Aiken, S. C., and although a rail haul of ninety miles was involved there were, at the time, factors favorable to the Harleyville location offsetting, in part, the cost of transportation of clay, There are, no doubt, locations in the general area where clay and limestone deposits may be found in closer proximity hut they had not been proven up when the project was undertaken. Availability of power from the Santee-Copper development whose high tension line skirted the Property; ready access to rail and highway transportation; the promise, realized in actuality, of ample supply of soft water; and availability of labor supply were other factors of importance. Moreover, the general area has the decided advantage, for this type of operation, of a mild climate. Construction was begun in August 1943, and although stopped for a time priding review of the general situation that fall, it was resumed, albeit with less urgency, and substantially completed by the end of 1944. The total capital cost was approximately $3,460,000 The first alumina was turned out in April 1945 and from then until the plant was closed down in July 1946, the operations followed a succession of periods of production alternating with periods in which minor alterations were made in conformity with experience. Steady improvement was made and at the time Operations were discontinued it was considered that experience and the necessary engineering were in shape for removal of the remaining bottlenecks and a somewhat major revision in one section of the plant. Unfortunately this could not be carried out and it was necessary to discontinue the

Jan 1, 1949

By F. L. Fayers, J. W. Sheldon

It has been a matter of concern to the petroleum industry to determine what effect the capillary term has on saturation profiles, since these profiles determine ultimate economic oil recovery. In their original solution of the non-capillary two-phase flow problem, Buckley and Leverett&apos; showed that solutions of the two-phase flow equations became multiple-valued in saturation. Since it is physically unrealistic for saturation to have more than one value at a given position, Buckley and Leverett resolved this difficulty by introducing a saturation discontinuity or "shock". They evaluated the strength and position of the shock from material balance considerations. Recently Sheldon, Zondek and Cardwell&apos; have demonstrated that the method of characteristics can be used to solve this and related problems. Owing to capillarity it is physically unlikely that saturation shocks occur in an oil reservoir. Buckley and Leverett suggested that the effects of interfacial tensions should be included in the fluid flow equations, if the solutions are to give continuous single-valued saturation distributions. An analogy to this problem is encountered in the theory of wpersonic compressible flow where the introduction of the viscous term eliminates shock behavior. However, the mathematical form of the capillary term in the porous medium fluid flow equations differs from the viscosity term in the Navier-Stokes equation. This paper demons~rates that capillarity does eliminate the triple-valued behavior and indicates the magnitude of capillary and gravity effects on the saturation profile for a variety of systems. DERIVATION OF LAGRANGIAN AND EULERIAN FORMS OF THE DISPLACEMENT EQUATION The equations representing capillary, two-phase flow are given by applying Darcy&apos;s law and a material balance condition for each phase. The familiar Darcy and material balance equations are as follows. ^U*^0......<3, ^^r=°......c« S is the saturation of the water phase, and B the inclination of the porous medium. Pressures in the oil and water phases differ by the capillary pressure associated with the interfacial tensions, P.-P.=pAS)..........(5) where p, is determined experimentally. The porous medium is said to be oil-wet if P, > Po and water-wet if P, < Po. If the fluids are incompressible and the total volumetric flow rate, q, is constant, then v, f v,, = q...........(6) By eliminating P.. /»,, and v, from Eqs. 1, 2, 5 and 6 it follows that It is convenient to normalize the variables in the problem after the manner of Handy and Hadley% so that Eqs. 3 and 7 become where j,, = —, the fractional flow of water; (F, is a capillary pressure normalizing constant.) Substitution of Eq. 9 into Eq. 8 gives In this partial differential equation for an initial value

By W. I. Garms, J. L. Stevens

BECAUSE of the many conflicting opinions regarding the functioning of the ball load and the characteristics of ball wear in ball-mill grinding, the following resume of reports covering recent tests conducted at the Hayden plant of the Kennecott Copper Corporation is submitted with the view of pointing out certain salient features that should throw additional light on the subject. MILLS USED Cylindrical drum mills with trunnion feed and discharge were used. The discharge trunnion was I3 1/2 in. in diameter. Dimensions inside the lining were 6 ft. 6 in. in diameter and 9 ft. 6 in. long. The mills were Britannia-lined and nine lifter bars projected 2 in. from the lining. The mill speed was 16 2/3 r.p.m. Feed to the mill was 9 per cent on 10-mesh, 85 per cent on 100-mesh, 77 per cent solids. New balls measured 2 in. in diameter and the approximate weight of the ball load was 40,000 pounds. INITIAL TEST During a test that was being conducted to determine the relative economic value of standard cast balls vs. forged steel balls in the local grinding operation, certain observations were made in regard to the functioning of the various ball sizes and the characteristics of ball wear, which warranted a supplementary test for the purpose of confirming such observations. For the initial test, the necessity for developing a graduated load of the new forged balls before the test could be started was realized, but the time required to obtain a graduated load and the functioning of the various ball sizes was not clear. No advantage could be seen for starting out with a full load of forged balls, as this initial load, at the least, would have to be completely ground out and replaced by daily additions of balls before the graduation could be complete and a normal static condition of ball sizes reached. For a considerable period prior to exhaustion of this initial charge there would be an unusual quantity of worn balls of a certain size, which would not be typical of a normal charge and could influence the characteristics and functioning of the entire load. Therefore, it was decided to start by daily addition of the forged balls to a mill that contained a normal operating load of cast balls. After 88 days of operation the mill was shut down for inspection. By that time the load consisted of a mass of forged balls functioning around a large core of the residual material of the former cast-ball load. This core of cast slugs, locally referred to as "the kidney," was well defined from the forged-ball load.

Jan 1, 1946

By R. P. Smith

The carbon content of a number of alloys ranging (before carburitation) from purified iron to purified cobalt was determined for fixed activities of carbon by equilibration with each of six CO-CO2 atmosphere at 1000°C. At fixed carbon activity, the carbon content decreases continuously with increasing cobalt content. For concentrations of 0 to 50 at. pct Co the results can be expressed by a linear interaction equation of the type In y2 = 7.76n2/(n1+n3)+ 2.30 n3/(n1+n3) where y2 is the activity coefficient of carbon (standard state such that y2— 1 as n2 — 0 in the Fe-C binary) and n1, n2, and n3 are the number of atoms of iron, carbon, and cobalt per unit weight of alloy. The results are also given in terms of the a functions. In the same concentration range, the results are in accord with the regular solution theory of Alcock and Richardson. THE carbon content of Fe-Si, Fe-Mn, and Fe-Ni alloys at fixed activities of carbon has been reported previously.&apos;-, The present work extends R. P. SMITH, Member AIME, is Research Chemist, E. C. Bain Laboratory for Fundamental Research, United States Steel Corp., Research Center, Monroeville, Pa. these measurements to a series of Fe-Co alloys. These alloys, like the Fe-Ni alloys, are fcc ("aus-tenitic") over the entire range of composition at 1000°C. EXPERIMENTAL PROCEDURE The alloys were made from electrolytic cobalt and Plast-Iron in a small vacuum-melting furnace. A small amount of graphite was added to each melt to keep the oxygen content low. The ingots were analyzed for iron and carbon: the amount of cobalt was obtained by difference. Analysis of the top and bottom of the ingots indicated that they were uniform in alloy content. Except for the pure cobalt the ingots were hot-rolled to approximately 0.15 cm thickness, cleaned by sandblasting, then cold-rolled to about 0.045 cm. We were not successful in cold rolling the pure cobalt; therefore it was hot-rolled to the final thickness. The last traces of surface oxide were removed by acid pickel, then the samples were treated in a hydrogen-2 pct water vapor mixture at 1100°C to remove carbon and most of the oxygen. The carburizing procedure was similar to that previously described.1,2,4 The alloys in the form of strips about 1.7 cm wide, 7.5 cm long, and 0.04 cm thick, together with a similar specimen of purified iron, were equilibrated with each of six CO-CO2 mixtures of fixed composition, then quenched in the

Jan 1, 1965

By C. P. Proctor, E. D. Hummer, J. A. Younkins

COMPARATIVELY little has been published relating practical experience with the control of the solid content of washery water systems that must be "closed." A "closed" system is one that for some reason may not be cleaned by the emission of any effluent of high solid content to be replaced by clean water in order to hold the solid content of the waste water within reasonable bounds for efficient washing. It is the purpose of this paper to describe practical experience with such a water system and to present data that it is hoped may be of some use to those who are faced with a similar problem. The washery clarification system described comprises three thickeners and three filters, and although it is admitted that each of the fields of water settling and thickening and filtration offers a wide scope for investigation, the authors are not attempting to present the answers to these problems but to set down the results of some experience along these lines, making no claims as to the ultimate solution of the problems in hand. The subject of water clarification has vast ramifications; for example, the science of flocculation as applied to water settling and filtering opens up fields of wide possibilities of which as yet little is known. A comparatively recent entry into the field of water treatment, which may offer possibilities of practical application, is the use of "wetting" agents such as "Aerosol." GENERAL ARRANGEMENT OF WATER SYSTEM So that the reader may obtain a comprehensive picture of the factors concerned in the control of a closed water system, it is necessary to give a brief description of the washing plant that is serviced by the water system to be described in the following pages. As we are not primarily concerned with the washing plant in detail, only a brief description of the flow of coal as it relates to the water system will be given here. For a more detailed account of this type of washer, the reader is referred to earlier papers.1.2 The washer, which handles 0 to 4-in. raw coal (750 tons per hour) is an American Rheolaveur plant of two complements; the coarse coal or sealed discharge plant and the fine coal or free discharge plant. The sealed discharge plant (Fig. r) consists of twin 48-in. primary launders, each having two washing boxes, supplemented by one 32-in. rewash launder, which also has two washing boxes. "Push" water for washing and conveying coal in these launders is supplied from a constant-head tank under pressure head of 10 ½ ft. at the point of introduction at the feed end of the launder. This tank also supplies water for upward or vertical currents used in the Rheolaveur or washing boxes. Each box discharges into a sealed boot elevator (Fig. I).

Jan 1, 1942

By P. Nyul, S. Caplan, M. F. Lamorte, T. Gonda

Therrnal resistance is measured on GaAs laser diodes in the temperature range 77" to 300°K. These data show that typically the thermal resistance increases fifteen times from 77 to 300°K. The increase is attributed to the decrease in thermal conductivity of Gds. Nornzalized thermal-resistance experimental and calculated from published thermal-conductivity data) values are compared and good agreement shown. In many practical applications, it is desirable to obtain high average power output from laser diodes. Average power output is determined in large measure by the external efficiency and thermal resistance of the diode.&apos;-3 The average input power with which a laser diode may be pumped is dependent on the thermal resistance. Thermal resistance has been studied over the temperature range from 77" to 300°K which is usually of interest in practical applications. This correspondence presents data on thermal resistance for GaAs laser diodes and compares them to values calculated from thermal-conductivity data. Junction heating arises from three distinct components: 1) joule heating, 2) nonradiative recombinations, and 3) reabsorption of radiation. In high-threshold lasers, joule heating predominates.4 Typically, the nonradiative recombination is not negligible and contributes significantly to junction heating. The loss caused by reabsorption may be of the order of the joule-heating loss.&apos; The fabrication and the geometry of laser diodes have been described in other publications.5 The data presented were obtained from diodes which were fabricated from the same region of a wafer to minimize differences. Fig. 1 shows four types of heat sinks employed in this study. In types 1 and 2, the laser cavities are alloyed; pressure contacts are employed in types 3 and 4, and indium is plated on the inside faces to improve the intimacy of contact between the diode and the heat sink. The thermal resistance R is defined by the following relationship: where AT is the junction-temperature rise above the heat-sink temperature for the diode power dissipation Pd. For measurement of thermal resistance, a direct current is passed through the junction, in the range from 1 to 50 pct of threshold. The diode is placed in the laser mode by superimposing a current pulse of low-duty cycle on the direct current.3 Junction-temperature rise is obtained from the time-resolved laser spectrum at the leading edge of the current pulse.&apos; Typically, the laser line shift is 2.5A per OK. The resolution employed in these experiments is 0.5A; therefore, the accuracy in the temperature measurement is 0.2K. Typically, the temperature rise is 2" to 20°K. The power dissipation is the difference between the dc power input, and the radiation power exiting the diode. The measurement accuracy is less than 1

Jan 1, 1968

By Robert P. Burrows

Discussion of the paper of R. P. Burrows, presented `at the New York meeting, February, 1916, and printed in Bulletin No. 107, November, 1915, pp. 2237 to 2245. EDWIN M. CHANCE, Wilkes-Barre, Pa.-I have been very much interested in the comprehensive and able discussion of the important topic of mine lighting just given by Mr. Burrows, especially as I know by common report Mr. Burrows&apos; ability and his close touch with the development of the electric mine lamp, and its use in mines in connection with a portable electric lamp. It is regrettable that Mr: Burrows was not able to acquaint himself more closely with the cost of mine illumination, the actual cost, as produced by a portable light as carried by miners, whether a flame safety lamp or an open light. Had he been able to acquaint himself more closely with these data, he perhaps would not have arrived at some of the conclusions which he has presented to us in his paper. For example, we find Mr. Burrows advises that the open torch is by far the cheapest source of illumination, as produced by a portable light. I am very sure that is not the fact. The cost of miners&apos; oil varies from about 25.c. a gallon to 90b. a gallon. The various States have passed laws from time to time rendering it mandatory that the miner shall use, whether he is willing or not, a so-called better or higher grade of burning oil. These laws have entailed the use of an oil so expensive that the use of the oil torch has become practically impossible, and the acetylene or carbide miners&apos; lamp has largely taken its place. I have evolved a few figures showing the relative efficiency and usefulness of the open oil torch and the acetylene miners&apos; lamp. For example, the acetylene miners&apos; lamp consumes or burns 4 liters of acetylene per hour, or 0.14 cu. ft.; the oxygen consumed is 10 liters, or 0.35 cu. ft. per hour; air consumed, 50 liters, or 1.76 cu. ft. per hour; black-damp produced, 40 liters, or 1.41 cu. ft. per hour; air rendered extinctive to the acetylene flame, 112 liters, or 3.95 cu. ft. per hour; carbide con- sumed, 13 g. per hour. The open oil torch, on the other hand, burns about 20 g. of oil per hour; the oxygen consumed is 48 liters, or 1.69 cu. ft. per. hour; air consumed, 240 liters, or 8.47 cu. ft. per hour; blackdamp produced, 191 liters, or 6.73 cu. ft. per hour; air rendered extinctive to oil flame is 1,413 liters, or 49.92 cu. ft. per hour.

Jan 5, 1916

By John Chipman, Raymond L. Orr

Recently reported high -temperatlnre thermal data have been incorporated in a reuieu&apos; of the therinody-r~ar,zic propesties of iron. The recent data permit mot corlsisterlt and reliuble choices to be made for many of the pvoperties, resnlling in bettev eslablished tabrtlations of the thermodynamic functions. Values of Cp, h°t- - H°298. s°t - S2sa, and (G°T -H298/T are tabulated for -the solid and liquid phases of iron in both stable and melastable regions. Consistency of the tabulations has been maintained to the precision necessr~vg to yield the Gibbs energy change belween the bcc (a, 8) and fcc (y) phases of iron at temperatures pertinenl to alloy studies. THE importance of the metal iron has stimulated many investigations of its thermodynamic properties at elevated temperatures. Reviews and summaries of the data have appeared from time to time. Among these the 1951 paper of Darken and smith&apos; deserves highest recognition for the quality of its criticism and for the precision and thermodynamic consistency of its tabulations. These and the more recent table of Kaufman, Clougherty, and weiss2 reproduce the Gibbs energy of the (a, 6)--Y transformation with sufficient sensitivity for alloy studies. Other recent compilations of equal validity3-&apos; make no attempt to do this. Since this Gibbs energy change is not greater than 22 cal per g-atom over the temperature range 1100° to 1800°K, it is obviously desirable to tabulate values of (G°t - H298)/T to four places of decimals even though the absolute accuracy would not warrant such precision. The necessity for reconsideration of the problem is emphasized by the recent publication of several important researches on the enthalpy and heat capacity of iron. The older work is well-covered in the reviews cited and will be mentioned here only when actually employed in the tabulations. GAMMA IRON Transition temperatures are taken from the selec -tions of Elliott and Gleiser,5 based on the work of Boulanger7 and earlier workers. For the a-y (A3) transition at 1184°K we retain the value 215 cal per g-atom adopted by Darken and smith.&apos; It has received confirmation in the recent work of Dench and Kubaschewski,&apos; and of Braun8 using adiabatic calori-metry, and of Wallace, Sidles, and Danielson10 using a pulse heating technique. The enthalpy data of Olette and ferrier" and of Anderson and Hultgren12 at higher and lower temperatures may be extended to 1184°K to give the following values which are adjusted to the required value of ? 1184 Ha,1184 Ha 298 = 8030 cal per g-atom [11 hr,1184-ha298 = 8245 cal per g-atom [2]

Jan 1, 1968

By W. H. Somerton

The effects of drilling variable on rotary drilling rates and efficiencies have been studied by a series of laboratory drilling tests. TWO-cone 1.25-in. diameter hits were used to drill vertically upwards into rock samples at corltrolled weights and rates of rotation. Shale, sandstone and specially prepared concrete samples were used in this study. Power input to the ririllirig systeln was measured and drilling chips collected for energy — .size reduction studies. Reasouably good correlations between drilling variables and rates of penetrotion were found. Qrrantities that are difficult to evaluate include rock strength parameters and the effects of bit wear. Effects of bit size and geormetry require further investigation. Analysis of the drilling chips confirmed the premise that, for rocks containing two or more mineral constituents of diflerent strengths, a greater amourrt of rock breakage occurs in the weaker constituent. Drilling conditions which required greater amounts of energy produced finer drilling chips. As bit tooth wear progressed, drilling chips became finer. Efficiency of rotary drilling as a rock breakage mechanism was extremely low. Comparison was made with theoretical energy requirements and with energy requirements for size reduction by comminution methods. INTRODUCTION Many technological developments and innovations have made possible the successful drilling of oil wells by the rotary method to depths exceeding 20,000 ft. Bigger, more powerful rigs, better steels, improved bit design and more careful control of the circulating system have all contributed to this success. Despite these advances comparatively little is known regarding the basic mechanism of rock breakage by the rotary drilling process. Future improvements and, in particular, reduction of drilling costs will undoubtedly require a clearer understanding of the variables controlling effectiveness and efficiency of rotary drilling. Rotary drilling is inherently an inefficient means of producing rock breakage. Loss of energy in transmission from the surface to the cutting mechanism may be large, thus limiting the total amount of energy which may be applied to rock breakage. Increased efficiency may be realized by improved energy transmission methods such as the turbodrill. Conversion of transmitted energy to rock breakage by use of rotating cutter teeth is likewise inefficient. Improvement of cutter efficiency has been obtained essentially by empirical means —changes in bit tooth size, shape and spacing, and tooth deletion to give clean bottom-hole patterns. The present work was undertaken to investigate the factors controlling rates of bit penetration under laboratory drilling conditions. Laboratory drilling apparatus is shown in Fig. 1. The effects of rock strength and bit wear on drilling rates were investigated. Drill cuttings were analyzed to determine the character of breakage and to compare this with other methods of producing rock breakage. VARIABLES INVOLVED IN ROTARY DRILLING Some of the variables controlling bit penetration rates can be studied conveniently by dimensional grouping. R/DN f(F/D2S) or

By R. R. Knuckey

HAVING been associated with the operation of the de Vooys process for coal, which has treated 13,000,000 tons per annum, and recognizing the process as of value in ore sorting, Huntington, Heberlein and Co. Ltd. (Great Britain) in 1936 made an examination of all other existing methods. In none of the pioneer applications had recognition been given to the necessity for maintaining certain conditions that are essential for sorting certain ores if a truly efficient separation by gravity alone is to be assured. Consequently, in 1936-1937 Huntington, Heberlein and Co. Ltd., with Pearson, employing certain ideas connected with the de Vooys application, commenced the development of a separator that would allow conformance with optimum conditions in the zone where separation is effected. Later, Holt and Wade added to the value of the original application by Wuensch, Rakowsky and others, whose conceptions are being successfully operated today. Meantime, Trostler had enhanced the value of the original H-H. process by improvements that are now embodied in operating plants, and today the process incorporates an improved apparatus and a stable medium that produce the optimum conditions for separation. These conditions are: 1. Elimination of disturbing currents in the separating zone; 2. Reduction to a minimum of medium-density differences throughout. the separating zone; 3. Stable medium of minimum viscosity to permit the unretarded fall of mineralized particles; 4. Removal of products of separation with minimum disturbance to the separating zone. DESIGN OF SEPARATOR The H-H. sink-and-float separator consists of a pyramidal tank in which the separation actually takes place, connected at the bottom with an inclined bucket elevator. The tank and the elevator are filled with medium to within a few inches of the top of the tank. Medium is fed into the separator at a variety of points, according to the type of ore being treated, and it is withdrawn from the separator at two points: (1) together with the float product, by means of a variable-speed paddle, at the surface of the medium, and (2) by flowing over an adjustable weir cut in the side of the elevator casing. The prepared ore is fed onto the surface of the medium. The intermediate paddles, which revolve at a slow speed of 2 to 5 r.p.m., move the floating ore across the bath from feed end to float-discharge end. The ore, which is heavier than the medium, sinks below the sweep of the paddles and reaches the boot of the bucket elevator, which discharges it without disturbance of the separating zone. These arrangements for withdrawing medium from the separator allow the use of minimum flows, which connote minimum

Jan 1, 1943

By R. R. Knuckey

Having been associated with the operation of the de Vooys process for coal, which has treated 13,000,000 tons per annum, and recognizing the process as of value in ore sorting, Huntington, Heberlein and Co. Ltd. (Great Britain) in 1936 made an examination of all other existing methods. In none of the pioneer applications had recognition been given to the necessity for maintaining certain conditions that are essential for sorting certain ores if a truly efficient separation by gravity alone is to be assured. Consequently, in 1936-1937 Huntington, Heberlein and Co. Ltd., with Pearson, employing certain ideas connected with the de Vooys application, commenced the development of a separator that would allow conformance with optimum conditions in the zone where separation is effected. Later, Holt and Wade added to the value of the original application by Wuensch, Rakowsky and others, whose conceptions are being successfully operated today. Meantime, Trostler had enhanced the value of the original H-H. process by improvements that are now embodied in operating plants, and today the process incorporates an improved apparatus and a stable medium that produce the optimum conditions for separation. These conditions are: 1. Elimination of disturbing currents in the separating zone; 2. Reduction to a minimum of medium-density differences throughout the separating zone; 3. Stable medium of minimum viscosity to permit the unretarded fall of mineralized particles; 4. Removal of products of separation with minimum disturbance to the separating zone. Design of Separator The H-H. sink-and-float separator consists of a pyramidal tank in which the separation actually takes place, connected at the bottom with an inclined bucket elevator. The tank and the elevator are filled with medium to within a few inches of the top of the tank. Medium is fed into the separator at a variety of points, according to the type of ore being treated, and it is withdrawn from the separator at two points: (1) together with the float product, by means of a variable-speed addle, at the surface of the medium, and (2) by flowing over an adjustable weir cut in the side of the elevator casing. The prepared ore is fed onto the surface of the medium. The intermediate paddles, which revolve at a slow speed of 2 to 5 r.p.m., move the floating ore across the bath from feed end to float-discharge end. The ore, which is heavier than the medium, sinks below the sweep of the addles and reaches the boot of the bucket elevator, which discharges it without disturbance of the separating zone. These arrangements for withdrawing medium from the separator allow the use of minimum flows, which connote minimum

Jan 1, 1947

By R. R. Knuckey

Having been associated with the operation of the de Vooys process for coal, which has treated 13,000,000 tons per annum, and recognizing the process as of value in ore sorting, Huntington, Heberlein and Co. Ltd. (Great Britain) in 1936 made an examination of all other existing methods. In none of the pioneer applications had recognition been given to the necessity for maintaining certain conditions that are essential for sorting certain ores if a truly efficient separation by gravity alone is to be assured. Consequently, in 1936-1937 Huntington, Heberlein and Co. Ltd., with Pearson, employing certain ideas connected with the de Vooys application, commenced the development of a separator that would allow conformance with optimum conditions in the zone where separation is effected. Later, Holt and Wade added to the value of the original application by Wuensch, Rakowsky and others, whose conceptions are being successfully operated today. Meantime, Trostler had enhanced the value of the original H-H. process by improvements that are now embodied in operating plants, and today the process incorporates an improved apparatus and a stable medium that produce the optimum conditions for separation. These conditions are: 1. Elimination of disturbing currents in the separating zone; 2. Reduction to a minimum of medium-density differences throughout the separating zone; 3. Stable medium of minimum viscosity to permit the unretarded fall of mineralized particles; 4. Removal of products of separation with minimum disturbance to the separating zone. Design of Separator The H-H. sink-and-float separator consists of a pyramidal tank in which the separation actually takes place, connected at the bottom with an inclined bucket elevator. The tank and the elevator are filled with medium to within a few inches of the top of the tank. Medium is fed into the separator at a variety of points, according to the type of ore being treated, and it is withdrawn from the separator at two points: (1) together with the float product, by means of a variable-speed addle, at the surface of the medium, and (2) by flowing over an adjustable weir cut in the side of the elevator casing. The prepared ore is fed onto the surface of the medium. The intermediate paddles, which revolve at a slow speed of 2 to 5 r.p.m., move the floating ore across the bath from feed end to float-discharge end. The ore, which is heavier than the medium, sinks below the sweep of the addles and reaches the boot of the bucket elevator, which discharges it without disturbance of the separating zone. These arrangements for withdrawing medium from the separator allow the use of minimum flows, which connote minimum

Jan 1, 1947