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By G. F. Loughlin, C. H. Behre

THE DEVELOPMENT OF CLASSIFICATION WHAT is the use of a classification of mineral deposits? From the days of Agricola, the founder of the science of ore deposits, successive authors on the subject have attempted classifications, none of which have attained unanimous indorsement of miners, engineers, or geologists. The miner and engineer naturally have little regard for a classification that does not help them to find ore. Mining geologists may share this feeling, and deplore the overemphasis that may be placed on classification by their academic brothers. Apparent effects of such overemphasis may be reflected by students or new graduates who have grown to think that each deposit must fit perfectly into some pigeonhole of the more or less arbitrary scheme that has been drilled into them. Diversity in mineral deposits as in other natural phenomena, however, is too great to be fully expressed in the usual tabular classification, and experience is essential for the appreciation of the value or usefulness of a classification. The succession of classifications has closely followed the growth of the science of mineral deposits and reflects that growth, as well perhaps as the limited or comprehensive viewpoint of each proponent. The earliest classifications were based on metal content, and were superseded by those based mainly on form of deposit. Source of materials also became a basis, and was followed by physical and chemical conditions and processes. In spite of this development it is interesting to note that in many of the latest reports on mining regions the authors adopt special classifications suited for their particular districts. Some of these, like the earliest classifications, are based primarily on metal content, and others are most conveniently based on form, and it may be convenient to modify these by reference to source of material or temperature of deposition. Some, including the writers, after noting these different usages, have refused to worry much over classification as an ultimate objective, but,

Jan 1, 1933

By M. C. Chang, J. Dasher

The Crucible mine in Pennsylvania, operating on Pittsburgh seam coal, is rated at 5000 tpd. The washing plant, built in 1943, is rated at about 400 tph, using hydroseparator boxes to wash the coarse coal (4 x 5/16 in.) and hydrotator units for the fine coal (5/16 x 0). It is generally acknowledged that gravity concentration equipment cannot be used efficiently to bene-ficiate coal finer than about 65 mesh. Before the present flotation circuit was installed, Crucible recovered its fine coal slurry, predominantly —24 mesh material, by passing it through a battery of five 14-in. and one hundred and ten 3-in. Hey1 & Patterson cyclones in parallel with a 75-ft Dorr thickener. About two thirds went to the cyclones and one third to the thickener. Underflows from cyclones and thickener were filtered, respectively, in one 12-disk Eimco and two 4-disk Oliver filters fitted with stainless steel cloth. The resulting cakes were added to the coarse metallurgical coal. The filtrates, overflow from the 3-in. cyclones, and thickener overflow were recycled to the coarse coal plant. The three major problems encountered in this system became acute when the mine switched from track-mounted machinery to the more economical off-track equipment. The quality and uniformity of product declined, too much slime accumulated in the circuit, and waste pond disposal became highly impracticable. Lowering of Quality and Uniformity: Since the filter cake was not beneficiated in any way, addition of this material, which contained about 18 pct ash. would certainly lower the quality of metallurgical coal. Difficulties of handling the high-ash, high-moisture cake were compounded by the lack of adequate blending facilities at the mine, making it impossible to maintain uniform feed for the coke ovens. For a two-month test period, to verify the source of trouble, all filter cake was excluded from the metallurgical coal. The excluded filter cake was stocked in the yard. Slime Buildup: The moisture retained in the coarse coal products and the filter cake provided daily the only means of slime bleed from the circuit. As this bleed was not enough to offset the incoming fines in the raw coal feed, there was excessive buildup of slimes. To operate at all, it was necessary to purge the system each weekend by pumping all the solids from the thickener to a waste pond. The slimes built up during the week progressively increased the apparent viscosity and density of the circulating plant water, so that hydroseparators and hydrotator units had to be constantly adjusted. Uniformity, yield, and quality of the metallurgical coal suffered in consequence, and the filter cake ash and moisture contents increased from Monday to Friday. Limited Waste Disposal: Location of the plant does not permit unlimited waste pond area. Maintenance of a waste pond for the purpose of emptying the thickener during weekends became a costly burden that could not be tolerated. HOW THE PROBLEMS WERE SOLVED The flowsheet (Fig. 1) chosen to overcome these difficulties is considered the one most economical to install and operate. In a preliminary change made before the flotation circuit was installed, the cyclones and thickener were first operated in series, the overflow from the 3-in. cyclones, after flocculation, being discharged into the thickener and the thick-

Jan 1, 1960

By A. E. Lee

THE Blackwell Zinc Co., Inc., a subsidiary of the American Metal Co., Ltd., operates a horizontal retort zinc smelter at Blackwell, Okla. The plant has 14 furnace blocks of 800 retorts each, fired with natural gas on a 48 hr cycle. Over 13,000 tons of zinc-bearing material, chiefly sulphide flotation concentrates, are treated monthly to produce slab zinc and high lead-cadmium fume. In 1942 a program of rebuilding and modernizing the smelter was started. By 1947 furnace smelting capacity had been increased to a point where roasting and sintering facilities were inadequate, and it was necessary to purchase oxidized materials to supplement sinter production. The seven 210 ft Ropp roasters and three 42 in. x 44 ft Dwight-Lloyd machines then in use had been in service at least 20 years and were in need of major rebuilding. Thus it was entirely practical to consider all new equipment and a change of method rather than rebuilding and repairing obsolete units. A study of the problem indicated that roasting as such could be eliminated and roasting and sintering accomplished in one step by a modification of the Robson process,' which had been used since the early 1930's by the National Smelting Co., Ltd., at their plants at Avonmouth, England, and Swansea Vale, South Wales. Francis P. Sinn, General Manager, Zinc Smelting Operations, The American Metal Co., Ltd., who was familiar with the practice in England, suggested the use of one large machine for the entire operation from concentrate to sinter. One step sintering appeared to best meet Blackwell's plant requirements and indicated substantial savings in labor, gas, coal, and repair costs. Choice of Machine Size The sinter machine size was set at 12x168 ft for a rated capacity of 540 tons per day. This tonnage, produced on a five day week, would meet the seven day requirements of the 14 furnace blocks. The one large machine was quoted at a lower cost than two or more 6 ft wide machines of similar total capacity. Further, the larger machine could be housed in a smaller structure and only one set of equipment for charge preparation and delivery and for disposal of sinter cake was needed. One machine on a five day week made possible a concentration of the skilled operating personnel and required less men than a plant including two or more machines and related equipment circuits. Fewer units of equipment meant less maintenance, and the two down days weekly allowed ample time to repair and, if necessary, to make up lost production. Experience had indicated better sintering quality and rates with larger masses of material, not only on wider machines, but also in deeper beds. The ratio of windbox perimeter to area for the 12x168 ft machine is 0.179, compared to 0.353 for a 6x102 ft machine and 0.617 for a 42 in. x 44 ft machine. This meant less air leakage with resulting fan power savings and less spoilage of charge along the pallet sides. Performance Initial operation of the new sinter plant was made in November 1951 and regular production attained late in December. The average product sinter output during 1952 and the first half of 1953 has been 18.2 tons per hr. The average for one month has been as high as 22.4 tons per hr. Considerable experimenting with varied operating conditions accounts in part for the below capacity — 24 tons per hr — average output, and work to further improve production rate continues. A typical sinter analyses is 66.0 pct Zn, 0.3 pct Pb, 0.1 pct Cd, 0.3 pct S, 8.0 pct Fe, 2.0 pct SiO,, 0.8 pct CaO, and 0.2 pct MgO. Use of this material has made possible increases in furnace burden and improved furnace operation over the former practice using sinter made from Ropp roasted concentrates. Better lead and cadmium elimination in sintering has permitted the furnace production of slab zinc lower in lead and cadmium. Anticipated economies of operation have largely been gained. The sinter plant is operated by seven men per 8 hr shift — one head operator, three equipment operators and three sweepers — plus one oiler on day shift only. While it has been necessary at times to operate seven days a week to produce the required sinter tonnage, the five day work week usually has been adequate. Consumption of natural gas for sinter bed ignition is 200,000 to 300,000 cu ft per day. Green Ore Sintering Practice The 30 to 31 pct sulphur content of the —200 mesh zinc concentrates is the fuel used to sinter the charge, no coal addition being required. In the feed to the machine, sufficient concentrates are added to crushed return sinter fines containing 0.3 to 0.5 pct sulphur to produce a charge averaging 5.0 to 6.5 pct sulphur. Since the return sinter used in Blackwell's practice is varied from — 1/2 to — 1/8 in., the actual sintering mixture of fine sinter and concentrates is somewhat higher in sulphur. The coarser sinter particles are too large to resinter and merely aid porosity in the sinter bed. The ratio of concentrates to return sinter in the charge ranges from about 1:4 to 1:5.5. Variations are based on the appearance of pried up bed sections, bed exit gas temperature trends, windbox suctions, and return sinter size. Sufficient sulphur must be used to obtain fritting of the charge into a soft sinter cake and to aid in the elimination of lead and cadmium. Excessive feed sulphur will result in partial slagging of the cake impairing porosity and prolonging sintering time.

Jan 1, 1954

By G. R. M. Del Giudice, A. M. Sadler, Arthur F. Taggart, M. Hassialis

FLOTATION of sulphide minerals and native metals is no longer a practical difficulty. The underlying scientific principles of the method, although not explored in anything like complete detail, have been formulated with sufficient breadth so that the direction of attack on any specific case is clearly indicated. It has been known since the earliest days of flotation that certain of the nonsulphide nonmetal minerals-e.g., sulphur and graphite-would respond to flotation by oils as sulphides do; therefore they have been grouped with sulphides in many patents. But since the development of the chemical theory of collection, 1,2,3 this similarity in action has been one of the facts cited in disproof of the theory by opponents of the hypothesis. 4 5-1 Air levitationt of certain oxidized minerals, both metalliferoust and nonmetalliferous, and separation thereby from other oxidized minerals, is a recent development in practice, although the discovery of the method dates back 17 years .7 This paper presents evidence that the principles underlying the air levitation of the nonsulphide nonmetal minerals are the same as those utilized in the flotation of the sulphides and native metals. It will be apparent immediately to anyone with knowledge of mineral concentration that separation of air-levitated minerals in water from minerals not so levitated is a simple matter of water-gravity concentration, the separating device employed depending solely upon the size of the particles treated and the buoyancy differences effected by the levitation. The usual separating means employed are: (1) a flotation machine for fine, slimy and for well aerated granular material; (2) shaking tables for granular less aerated material. By using large quantities of oil, it is possible to form substantially air-free oil-bound masses or granules of nonsiliceous minerals and to

Jan 1, 1937

By Y. C. Liu, H. Margolin

Investigation of martensite habit plane in water-quenched Ti-Mn alloys was carried out in the range of manganese contents between 4.35 and 5.25 pct. On the basis of 22 measurements, the poles were observed to fall into two groups, indicating the existence of two habit planes. The Miller indices of the poles close to these two groups were found to be <334>ß and <344>ß. TITANIUM has an allotropic transformation at 882.1° ± 0.5C1 by which the high temperature ß (body-centered cubic) phase changes to a (hexagonal close-packed) phase. It is well known, as in the case of low carbon steel, that the unalloyed high temperature ß phase cannot be retained no matter how rapid the cooling rate. However, the addition of alloying elements usually lowers the transformation temperature continuously and stabilizes the ß phase to a temperature much lower than the allotropic transformation temperature in pure titanium. A typical example of this behavior is found with manganese-alloying of titanium. Fig. 1 shows the titanium-rich Ti-Mn equilibrium diagram as established by Battelle Memorial Institute.&apos; The addition of manganese lowers the ß transus line and the eutectoid reaction takes place at 550°C, the eutectoid composition being 20 wt pct Mn. However, upon quenching to room temperature from the ß field, the ß phase can be completely retained with much less manganese, see Fig. 2. If the composition is less than that required to stabilize the ß phase completely, a plate-like structure, similar to that of martensite in other alloy systems, appears as shown in Fig. 3. Much lower alloying content merely intensifies the amounts of these transformed structures, as shown in Fig. 4. This paper presents the results of an investigation of the habit plane of these martensitic structures in Ti-Mn alloys. Experimental Procedure Investigation of the pole of a given platelet was made using the two-surface method of analysis." In order that a given martensite trace on two surfaces be readily recognized both at low and high magnifications, it is desirable that only a small amount of martensite be present. From a review of the data obtained at Battelle Memorial Institute4 on quench hardening in Ti-Mn alloys and the results of Duwez" on transformation in Ti-MO alloys, it was decided that the most appropriate Ti-Mn composition for this study would be about 6 pct Mn. A group of alloys containing from 3 to 10 pct Mn was made to bracket this composition. The choice of the 6 pct alloy was apparently proper, since it was later learned that approximately 5.5 to 6 pct Mn is required to stabilize the ß phase on water quenching.

Jan 1, 1954

By J. L. Rodda, R. L. Wilcox, E. C. Truesdale

In recent years various problems in connection with research work on the preparation and properties of zinc-base alloys have required reliable information concerning the constitution of the zinc-rich portion of the iron-zinc equilibrium diagram. On numerous occasions rather positive indications were obtained that, for example, the solubility of iron in zinc was incorrectly represented on the current equilibrium diagrams. A careful study of all the available literature relating to the iron-zinc system led to the conclusion that much of the early work was unsatisfactory as measured by present standards of experimental accuracy, and as a result it was decided to redetermine the location of the liquidus curve and of adjacent portions of the diagram. Part I of this paper comprises a description of the experimental methods and a summary of the results by means of which the zinc-rich end of- the iron-zinc diagram has been revised in the following ways: 1. Measurements of the solubility of iron in liquid zinc by several different sampling methods, leading to a relocation of the liquidus curve. 2. Microscopic, conductivity, and X-ray studies of the solid solubility of iron in zinc, which, together with thermal analyses, have resulted in new values for the solid solubility and the eutectic temperature and composition. 3. Heating and cooling curves, obtained by a refined differential method, leading to new locations for the two peritectic horizontals. In Part I1 of the paper the attempt is made to review all of the published work having a bearing on the constitution of the iron-zinc system, and to correlate the present results, and some recently published data involving the use of X-ray technique, with the earlier work, much of which was done from 25 to 30 years ago.

Jan 1, 1936

By H. D. Strain

Potash is a generic term used to describe a number of compounds containing the element potassium (K), which is one of the three major plant nutrients. Potash content of these compounds is commonly expressed as percentage potassium oxide (K20), and the market value of the products is usually dependent on contained K20. Most of the potash produced (approximately 95%) is used in agriculture as fertilizer. Potash products today are shipped by producing countries all over the world to meet the need for increased food production. WORLD POTASH RESERVES Apparent world potash reserves are shown in Table 15.2B.1 and total some 96 billion mt of recoverable potassium oxide. The Canadian reserves account for approximately one-half the known world reserves. The reserves shown represent the potassium oxide content of a variety of potash-bearing ores. The major ore is sylvinite (a mechanical mixture of KC1 and NaCl) from which KC1 or sylvite is recovered (63.2% K20 equivalent). Other ores, such as langbeinite (potassium, magnesium sulfate-22.7% K20) and carnallite (potassuim magnesuim chloride-17.0% K20) are also mined in various parts of the world, mainly for the potassium content. Theoretically, potassium is available in almost unlimited qualities from brines and saline waters. To date, the cost of recovery from salt brines has been greater than the costs by mining operations, but new technology could make brines and seawater an increasingly important source. POTASH PRODUCTION Potash production dates back to the 1860s in Germany, which was the only recorded producing country until 1912. By 1912 Germany was producing approximately 1 million tpy of K2O equivalent. Production began in France in 1913 and in the United States in 1915. World potash production and consumption are shown in Table 15.2B.2 for the years 1950-51 through 1967-68. In 1967 and 1968 the major potash-pro¬ducing countries in order of importance were those listed in Table 15.2B.3. By 1972 Canadian production had outstripped U.S. production. WORLD CONSUMPTION World consumption of potash for the years 1950-51 through 1967-68 is shown in Table 15.2B.2. Table 15.2B.4 shows total world consumption of potassium materials in 1967-68 by country. More than 70% of the world&apos;s potash is

Jan 1, 1976

By George Rice

THE A.I.M.E. Ground Movement and Subsidence Committee, pro-posed in 1920, held its first technical meeting in February 1923, under the able chairmanship of Mr. H. G. Moulton. The following list of papers and discussions, indicating the scope of the problems and relation to various conditions and method of mining, were given at that meeting by the distinguished mining engineers: J. Parke Channing, on Subsidence at Miami, Arizona; Howard N. Eavenson, on Mining an Upper Bitu-minous Seam after a Lower Seam Has Been Extracted; J. J. Rutledge, on Subsidence in Two Oklahoma Coal Mines; Benjamin F. Tillson, on Caving Cracks at the Franklin Mine, New Jersey; Dr. F. W. McNair, on the Dome Theory; William Kelly, on disadvantageous effects in Mining and Back-filling Working Upward a Steep-Fitching Iron-ore Bed; Eli T. Conner, on the good results of Back-Filling by Hydraulic Means in Preventing Destructive Subsidence in the Anthracite District; Thomas H. Claggett, on the merit of Systematic Pillar Drawing in Preventing Cracks in an Upper Seam; Henry H. Otto, successful results in the anthracite district in Mining Lower Seams before Upper Seams; Douglas Bunting, on Leading Factors Incident to Successful Mining of an Overlying Seam; Edwin Ludlow, on there being "less damage to buildings when there was complete removal of coal with complete packing than when room and pillar is used without extracting the pillars"; Edward O&apos;Toole on "bumps" in eastern Kentucky mines, which he thought were the result of geologic stresses: H. A. Buehler, on Upheaval of Surface in an Illinois Case in Advance of the Face; and R. Dawson Hall, on The Mechanics of Subsidence from Mining. The author of this paper discussed various kinds of problems of ground movement and subsidence in both metal mining and coal mining; the production of great landslides by workings near an outcrop, as at Turtle Mountain, Alberta, Canada; and general

Jan 1, 1939

By P. A. C. Wilson

Instantaneous outbursts of carbon dioxide in coal mines have occurred in Germany only in one part of the Waldenburg-Neurode mining district.&apos; This mining region comprises the northeastern fold of the "Innersudetian" coal basin. The basin (Fig. 1) is in the form of a geologic trough approximately 60 km. (37 miles) long and 30 to 35 km. (19 to 22 miles) broad and is enclosed on three sides by ancient truncated mountain ranges—the Riesengebirge, the Bober-Katzbachgebirge and the Eulenge-birge. Only the southeastern part of the trough is not enclosed by mountains of older rock formation. The floor of the basin before the building up of the coal measures was formed on the western, northwestern and northeastern borders by the material from Archean rocks (gneiss and mica schists); in the northern part over this was deposited the Silurian formation. The carboniferous period is represented both by the Lower ("kulm") and Upper Carboniferous, with an unconformity between the stages; that is, the kulm was tilted up and partly eroded before the Upper Carboniferous was deposited. The kulm is a salt-water formation (with marine fauna) but Upper Carboniferous rocks are nonsaline water deposits, with fresh-water flora only. The Lower Silesian coal measures were thus formed in a lime basin fed by the running water coming down from the mountains. As already indicated, the bottom series of the Upper Carboniferous formation, termed "Waldenburg strata" were laid down unconformably on the Lower Carboniferous. Over the Waldenburg strata there are the "Weisstein strata," an intermediate barren stage without coal beds, 400 to 500 m. (1300 to 1600 ft.) thick. Another unconformity followed, which is overlain by the &apos;(Saarbrucken" series, which contain the upper coal measures, and the (&apos;Ottweil" strata. In the last named, it is only on the southwestern fold that beds of coal are developed.

Jan 1, 1931

By Charles W. Tandy, Wm. I. Weisman

Sodium sulfate is an important industrial chemical, being one of perhaps a dozen or so chemical commodities that are produced and consumed in the United States in quantities exceeding one million short tons per year. In recent years over half of the country&apos;s production of sodium sulfate has come from natural sources, whereas 20 years ago only 25% was derived from these sources. Production of natural sodium sulfate from various types of deposits is the main source of this chemical in Canada and Mexico, and probably in Argentina, Chile, and the USSR. [Table 1] illustrates the growth in production of natural sodium sulfate in the United States. Mineralogy and Physical Properties Sodium sulfate is widespread in occurrence and is a common constituent of many mineral waters, as well as seawater. Many of the saline or alkali lakes throughout the world contain varying amounts of sodium sulfate. In its natural form it is found as the hydrous salt, mirabilite, commonly called Glauber&apos;s salt, and as thenardite, the anhydrous variety. The largest quantities occur in the form of mirabilite. Sodium sulfate is found in varying degrees of purity, from theoretically pure efflorescent crystals of mirabilite to combinations and admixtures of other salts and impurities. It is a common constituent of some brines, and it is from this source that much is extracted commercially. Sodium sulfate is also found combined with other compounds, such as the mineral glauberite, the double salt of anhydrous sodium and calcium sulfate, and blödite, the double salt of sodium sulfate and magnesium sulfate. While these and other minerals containing sodium sulfate are not economically important, they are of interest because of their frequent occurrence. The reader is referred to other publications (Cole, 1926; Dana, 1932; Grabau, 1920) for descriptions of these minerals. Only the two economically important minerals, mirabilite and thenardite, will be described herein. Mirabilite, Na2SO4 • 10H2O, contains 55.9% water of crystallization. It is noted for its efflorescence or spontaneous loss of water. On

Jan 1, 1975

By H. I. Aaronson, C. Laird

The kinetics of thickening and of lengthening of ?&apos; plates in an Al-3.93 pct Cu alloy in the temperature range 203" to 300" C were determined by means of transmission electron microscopy. The rate of thickening was found to be less than that allowed by volume diffusion control at all temperatures, by amounts which increased with decreasing temperature, in agreement with the predictions of a general theory of precipitate morphology.1 Thickening was treated on the basis of the ledge mechanism. Ledges were deduced to spread across the broad faces of ?&apos; plates at volume dzjrfusion-controlled rates, as anticipated from the disordered structure of their edges. Lengthening of 8&apos; plates, on the other hand, took Place more rapidly than allowed by volume dzjrfusion. This occurred despite clear morPhological evidence of a bmrier to growth at the edges of these plates. It was concluded that the misfit dislocation structure comprising the barrier requires that lengthening take place by a jog mechanism. The tnisfit dislocations, however, also serve as diffusion short circuits, and allow high overall lengthening rates to be achieved. In Part I&apos; it was shown that, within the range of aging temperatures and times studied, the broad faces of 8&apos; plates formed in Al-4 pct Cu are fully coherent with the a, matrix. Virtually .all of the dislocations present in these faces were found to have developed as a result of plastic deformation in the a phase. Such dislocations are thus "intruders", rather than the more usual misfit-compensating variety. The edges of 8&apos; plates were confirmed, by extension of the earlier studies of Mat-suura and Koda,3 to be made up of edge-type misfit dislocations, in sessile orientation with respect to lengthening of the plates. These interfacial structures should cause 8&apos; plates to thicken and to lengthen at rates less than those allowed under the condition of volume diffusion control, such as would be expected if the interphase boundaries had disordered structures.&apos; The narrow width of 8&apos; plates, the reproducible crystallography of their broad faces, and the appearance of these plates in cross section as octagons rather than as circular discs2 provide qualitative support for these deductions. The present study of the rate of thickening and the rate of lengthening of 8&apos; plates was undertaken in order to examine them on a quantitative basis. I) THICKENING KINETICS OF THE BROAD FACES OF?&apos; PLATES A) Literature Review. The measurements now available on the thickening kinetics of single-phase precipitate plates consist of one plot of the thickening of a proeutectoid ferrite plate in an Fe-C alloy,&apos; showing (as predicted) thickening rates less than those allowed by volume diffusion control. B) Experimental Procedure. Details of the preparation of the 4-3.93 pct Cu alloy used in this study have been previously reported.4 As in Part 1,&apos; transmission electron microscopy was the observational tool employed. A general description of the apparatus and procedures of the electron microscopy studies is given in Section I of Part I. In thin foils, 0&apos; plates tend to form at and parallel to the foil surface.&apos; A direct investigation of the thickening process by means of hot-stage transmission electron microscopy was therefore not feasible. It was thus necessary to use the conventional method of aging individual bulk specimens for a wide range of different times at the various temperatures studied. In each specimen, the thicknesses of a number of plates were measured. Since thin foils prepared from "bulk-aged" material contain a large proportion of grains with orientations near (001) , it was relatively easy to find, near the edges of the foils, the characteristic multifold patterns of intersecting extinction contours which indicate regions where the foil is exactly at an (001) orientation. The thicknesses of large numbers of plates were measured along the (200) branches of the "stars" so that the 8&apos; plates were precisely parallel to the optical axis of the microscope. Wherever possible, intersecting extinction contours were adjusted with the parameter s > 0 to improve the visibility of the plates in bright-field illumination. These precautions, in combination with taking the measurements at the thinnest parts of the foils, minimized the errors in the measurement of the thickness of the plates resulting from inexact parallelism to the electron beam. Since the plates were very thin, it was not easy to measure their thickness on the photographs. The techniques of enlargement and of microdensitometry were employed to minimize errors from this part of the measurement. A further source of possible error, that the plates can appear thicker because of contrast associated with mismatch normal to the plane of the plate, was also considered. The images of the plates were usually thinner than those of dislocations, however, and no anomalous changes in apparent plate thickness were observed when regions of foil containing plates were tilted through various diffracting conditions. Any error from this cause must therefore be small. Other sources contributing errors were: a) the microdensitometer traces per se and the subjective estimates of their peak limits, and b) slight fluctuations in magnification associated with small changes in the current of the objective lens of the electron microscope. The overall error probably amounted to no more than 5 to 10 pct. In order to obtain readily interpretable data on thickening kinetics, it is essential that the diffusion fields of adjacent ?&apos; plates not be allowed to overlap. Calculations&apos;-&apos; showed that this condition is definitely not ful-

Jan 1, 1969

ARTHUR KNAPP, Ardmore, Pa. (communication to the Secretary*). -The author has failed to consider the point which in my mind is the most convincing in advocating some sort of a bureau for the collection of well logs. I venture to say that 90 per cent. of the logs as taken today are absolutely worthless to any one but the one who has taken the log. This is largely due to the careless use of nomenclature. The author says on page 289, "He must ascertain which of the many factors might account for the dry hole in question and often only careful examination of well logs will convey this information." I believe that even a most careful examination of the average log reveals nothing. I have at hand a book on petroleum, by a well-known author. This book has a number of logs from fields all over the world, from which I shall illustrate. In a well log from Kansas I find "Soft gray shale" and in one from Humble I find "Hard blue shale." As a matter of fact the shale from Humble is softer than the shale from the Kansas well. The difference is that the drillers in Kansas, being used to limestone and

Jan 5, 1917

By A. W. Gauger

Coal as mined contains varying quantities of inorganic components (mineral matter) which, on combustion, produce the residue known as ash. It has long been realized that the weight of this residue does not equal the weight of the mineral matter originally in the coal, because of the chemical changes that take place during the ashing process. Moreover, the ordinary ash analysis presents the ash as a mixture of oxides with little or no clue to the identity of the mineral species originally present in the coal. It is only for pyritic sulfur that an attempt is made to determine a mineral component. So far as the authors are aware, there are practically no experimental data in the literature as to the percentages of the different mineral species present in coal. The oft-quoted table of Marson and Cobb1 cannot rightly be considered as data, since the authors merely state that the table "shows the probable forms in which these ash constituents exist in coal." In view of the attempts to calculate analyses and determinations of the calorific equivalent of coal to the mineral-matter-free basis, as well as the importance of the mineral species in clinkering, combustion and coking, it seems strange that so few data are available on this phase of coal technology. For example, numerous empirical corrections for use in the scientific classification of coal have been suggested to compensate for the losses from the mineral matter during ignition of the coal to ash. These include multiplication of the ash percentage by various factors such as l 1/10 and 156, the use of the Parr unit coal formula (1.08 A + 22/40 S, where S = total sulfur) as well as certain modifications of the Parr formula that attempt to correct for the organic sulfur present in the coal. Although many experiments have been made to determine the loss on ignition of the clays associated with the coal measures and correction

Jan 1, 1934

By Jonathan S. Jackson, Bayne B. McMillan, W. Joseph Schlitt

Various dump leach models have been developed by Kennecott, and these are reviewed with an eye toward production forecasting. Some of the models have been bated on first principles, utilizing the chemistry and physics of dump leaching. While providing considerable insight into the leaching process, these have proved to be too cumbersome for use in estimating production. Other models have been based on a declining leach rate for a single block of waste. These permit good short-term projections of solution grades and production, but are inflexible with regard to changes in mining parameters because the trends are based only on historical data. Therefore, a simple and flexible computer model has been developed to aid in leach production forecasting and to provide a tool for assessing the impact of different mining plans on leaching. This program, LEACH, is based on annualized blocks of waste; input parameters are described and examples of its use are given.

Jan 1, 1980

By H. I. Aaronson, H. A. Domian, G. M. Pound

Zener&apos;s two-parameter theory of the y a reaction in Fe-X alloys is extended to encornpass austenite-stabilizing as well as fewite-stabilizing elements, and is then cottzbitzed with statistical thermodynamic theories of Fe-C alloys to permit development of a ther-modynamic description of the proeutectoid fewite reaction in Fe-C-X alloys. Equilibrium tielines are calculated for tile a + y rep&apos;on from experimental data and frort extrapolations of the y/y +a equilibrium surface. Sectzovs of the y/y + 0 surface are calculaled for the tzetastable sstuation LYL u-kich alloyrtg eleitzents do not partition betzseen azrstenite andfer-rite. The finding that the metastable y/a + y curves Lie close to their equilibriuwz counterparts when X = Si, .Vo, Co, Al, and Cu but well below them when X = A/ln or Ari pr-elides a thermodunamic explanation for the expcri1?7e?ztally obserted absence of partition during the proeutec toid ferrite reartlon in the fortner alloys and for the occurrence of partition at hzgher ternperatures in the latter alloys. Cotarison ofno- patitzoa free-energy changes and y/y i a curr,es in Fe-C-X alloys with the equilibrium 11al1es of these quantities in Fe-C alloys furnishes additional qualitative insight into the irqluence of allojling elements upon the kinetics of the proeutectoicl ferrite reaction. THE introduction of a substitutional alloying element, X, appreciably complicates calculation of the thermodynamics of the formation of proeutectoid ferrite from austenite. As noted in the preceding paper,&apos; the positional entropy of the interstitial species is the principal component of the free energy of an interstitial solid solution with which theory has so far been able to deal. One would expect, however, that other components of the free energy of this type of solid solution may be significantly altered by the addition of a substitutional alloying element. Even if the assumptions are made, by analogy to the case of Fe-C alloys,&apos; that changes in the positional entropy of carbon represent a significant part of the thermodynamic effects of an alloying element, and that the remaining effects can be taken into account by fitting the equations developed on this basis to experimental data on Fe-C-X phase diagrams or on the activities of carbon in alloyed austenite and ferrite, the experimental information available on either of these quantities is not yet sufficiently accurate or extensive, respectively, to make such an approach useful. An attempt made by Zener on the former basis to explain the effects of alloying elements on the thermodynamics of Fe-C alloys in terms of a temperature-independent "free-energy change" (actually enthalpy hane&apos;) required to transfer 1 mole of an alloying element from austenite to ferrite, in which the "free-energy change" was determined by fitting the theoretical relationships developed to Fe-C-X phase diagrams, thus proved inadequate in part because of deficiencies in the available ternary-phase-diagram data.= Other difficulties of a more fundamental nature, however, also indicated the desirability of a different approach to the problem.3y4y6 zener6 subsequently proposed that the free-energy change associated with y — a transformation in pure iron can be decomposed into magnetic and nonmagnetic components. Alloying elements were assumed to affect those components separately. One parameter was used to describe the quantitative effect exerted on each component. These parameters were evaluated from Ms (martensite-start) temperature data and from other experimental information usually either readily available or measurable with acceptable accuracy. Zener applied this treatment only to the calculation of Fe-X phase diagrams in which X is a ferrite-stabilizing element. In the present study, this treatment is extended to include austenite-stabilizing elements, and then combined with treatments previously considered for Fe-C alloys1 to permit calculation of the thermodynamic quantities of interest in the austenite - proeutectoid ferrite transformation in Fe-C-X alloys, where a number of representative, and commonly used alloying elements are chosen for X, including Si, Mn, Co, Mo, Al, Cr, and Cu. The results are used to explain important features of the partition of alloying elements between austenite and proeutectoid ferrite, as reported in a companion paper7 on the basis of electron-probe analysis, and to provide some additional understanding of the influence of alloying elements upon the kinetics of nucleation and growth of proeutectoid ferrite. Portions of this treatment6 have been clarified in the course of a review by Kaufman and Cohen.3 The consolidated summary of these developments with which this section is begun provides a basis for further clarification of the treatment, from which extension to include austenite- as well as ferrite-stabilizing elements is a natural consequence. The division of the free-energy change associated with the 7 - a transformation in pure iron, into two independent components is formally stated as:

Jan 1, 1967

By A. Hansson, G. E. Tardiff

WHILE it is generally known that cold-worked low-and medium-carbon steels exhibit substantial increases in tensile rupture ductility with increased deformation velocity172 (up to the von Karman limit), the observations for annealed or normalized steel do not permit simple generalization. Various investigators, for example, have reported both small increases1,3 and small decreases4,5&apos; in rupture ductility with increased rates of deformation of normalized low-carbon steel. The purpose of the present note is to report on the dependence observed for strain-aged steel since preliminary work here indicates that strain aging results in the development of a consistent type of behavior that is understandable in the light of current knowledge of the lower yield dynamics for steels which deform inhomogeneously.6 This realization throws additional light on the influences responsible for the behavior of normalized steel. The results presented here were obtained for a 0.21 pct C, 0.012 pct N sheet (0.012 in. thick) steel which was received in a temper-rolled condition (approximately 3 pct reduction in thickness after process annealing); the as-received material had a spherodized structure with a ferritic grain size of ASTM 8. The material was selected for testing after 6 months room temperature aging, at which time the static lower yield and ultimate tensile strength measured 79 and 82 kpsi, respectively. ASTM 2-in. reduced section, 1/2-in.-wide specimens were used in smooth bar tests and 3/4-in.-wide specimens containing two 60-deg opposed edge notches (root radii = 0.002 in.) were used in notched bar tests. To facilitate strain measurement, the smooth bars were blued with ink and marked at 0.1-in. intervals along the gage length; the notched

Jan 1, 1968

By H. J. Levinstein, H. J. Guggenheim

RbFeF3 is a transparent ferromagnet with a large faraday rotation which permits the direct observation of magnetic domain structures in bulk crystals. If the position of dislocations within the crystal were known a study of domain wall dislocation interactions would be possible. In this paper an etch pitting solution is described which gives a one-one correspondent between the etch pits and the dislocations intersecting the (100) surfaces of RbFeF3 crystals. In addition, RbFeF3 is shown to have sufficient plasticity so that known arrays of dislocaticms can be put into the crystal. It is concluded that RbFeF, is an ideal material for the investigation of domain wall dislocation interactions. THE influence of lattice defects on the motion of domain walls in ferromagnetic materials has been studied extensively both theoretically and experimentally.4 The experimental data obtained has been qualitative in nature due to the lack either of a quantitative knowledge of the magnetic domain array in the material or of the nature of the dislocation configuration in the material. At room temperature RbFeFJ has the cubic perovskite structure. It is ferromagnetic below 86°K.5,6 It is also transparent and has a large faraday rotation in the magnetic state.7 The domain structure of this material is observable in three dimensions in crystals which are millimeters in thickness.&apos; Since RbFeF3 satisfies the requirements of having an easily visible three-dimensional magnetic domain structure, it was of interest to determine if the dislocation structure of the material could be determined and if the material was sufficiently plastic so that dislocations could be introduced into the as grown crystals. This is a report of the results of that investigation. EXPERIMENTAL PROCEDURE The crystals of RbFeFs used in this experiment were grown by a technique described elsewhere.5 In order to obtain flat strain-free surfaces, the samples were first cut to the crystal orientation desired and the surfaces were polished by employing standard metallographic techniques. The samples were then immersed in a solution of 1 to 3 g of KOH in 100 ml of water. This solution removes the damaged surface layers uniformly leaving a film on the surface of the sample. Etch pits are obtained by immersing the sample in an approximately 50-50 solution of 30 pct peroxide and concentrated acetic acid. The film formed during the KOH treatment turns deep red in color in the etching solution and flakes off the sample. The etching time required depends upon the pit size desired and

Jan 1, 1970

By C. J. Hicks

The labor policy of the Standard Oil Co. (New Jersey) is founded first of all on paying at least the prevailing scale of wages for similar work in the community; on the eight-hour day at the refinery, with time and one-half for overtime; one day&apos;s rest in seven; sanitary and up-to-date working conditions; just treatment assured each employee; opportunity for training and advancement; payment of accident benefits beyond the amount prescribed by the State compensation law; health supervision by a competent medical staff; payment of sickness benefits after one year&apos;s service; coöperation with employees in promoting thrift and better social and housing. conditions; and assurance of a generous annuity at the age of 65, guaranteed for life after 20 years of service. Most of these features have been a part of the company&apos;s policy for many years, but it is only during the past two years that the cooperation of employees in determining these matters has been definitely assured through industrial representation. Industrial representation, in the Standard Oil Co. (N. J.), is a principle rather than a procedure. It is built upon the belief that personal association of those interested in any problem leads to a mutual understanding and a fair decision as to what is right. Fully believing in this principle, representatives of employees and representatives of management evolved a simple plan, the basis of which is that it gives every individual employee representation at joint conferences on problems and fundamental principles affecting all those interested in the industry. It is based on cooperation, not antagonism; its operation makes perfectly clear both to management and to employees that their interests are identical, and not at variance with the interests of the stockholders, and that mutual understanding and cooperation insure progress and success for all. Furthermore, experience has definitely shown that representatives of the employees are not only alert for the employees&apos; interests but are as keen as the representatives of the management in determining and insisting upon fairness to the employer. Though the plan has been in operation nearly two years, it is an experiment, in that, being based on a principle rather than on cut-and-

Jan 1, 1921

By J. P. Morris

PREVIOUS investigations1,2 have shown that alloying elements in liquid iron influence the thermodynamic activity of sulphur and thereby affect the partition of sulphur between metal and slag in the desulphurization process. For example, the greater efficiency of desulphurization in the blast furnace as compared to the open hearth can be attributed in part to a higher level of sulphur activity in blast-furnace metal due to the higher concentration of carbon and silicon. In the present investigation, a short study was made of the influence of manganese on the activity of sulphur in liquid iron and iron-carbon alloys. In contrast to carbon and silicon, manganese was found to decrease the activity coefficient of sulphur; and in iron-carbon alloys it counteracts to some extent the influence of carbon. However, at manganese concentrations normally present in the blast furnace or open hearth, the effect of manganese is small. Since manganese sulphide has a limited solubility in iron, manganese can act, under certain conditions as a desulphurizing agent. Considerable data on the manganese-sulphur product in carbon-saturated melts were obtained in the investigation and have been included in this report. The experimental procedure was the same as that used in the earlier investigations on the effect of silicon&apos; and carbon&apos; on sulphur activity. Briefly, the method was as follows: The molten alloy, contained in a graphite or sintered alumina crucible, was brought to equilibrium at a constant temperature with a mixture of hydrogen and hydrogen sulphide of constant composition by bubbling the gas through the metal. Samples of&apos; the melt were taken for analysis at regular intervals by suction through a 2 to 3 mm bore silica tube dipped into the metal. The experiments were run in a graphite spiral resistance furnace with melts weighing 50 to 60 g. The gas bubbling tubes were made of sintered alumina and were 5/16 in. OD, 1/16 in. ID, and 24 in. long. Equilibrium was assumed to have been attained when the sulphur content of the liquid metal reached a constant value. During an experiment there was a rapid loss of manganese from the melt by volatilization. To offset this loss, small additions of manganese were made periodically. The rate of manganese addition needed to maintain a constant manganese concentration was determined in preliminary tests. In all of the experiments, deposits of manganese sulphide formed above the melts in a cooler region of the furnace. Apparently, these deposits resulted from a reaction between manganese vapor and hydrogen sulphide in the gas. To prove that manganese sulphide did not volatilize from the melts to a measurable extent, an experiment was run in which helium was bubbled through liquid iron containing both manganese and sulphur. Although manganese volatilized rapidly in this test, there was no appreciable loss of sulphur. Volatilization of manganese sulphide from a melt would have led to an apparent equilibrium condition in which the sulphur content of the metal was lower than the true equilibrium value. The experimental results are shown in the first seven columns of Table I. The data in the last two columns were obtained from the previous work on the effect of carbon&apos; and show what the results would have been in the absence of manganese but with temperature, gas composition, and carbon content of the metal remaining the same. Comparison of the last four columns show that, in the presence of manganese, the sulphur content of the metal increased at equilibrium and the activity coefficient of sulphur decreased. However, the results show that, for manganese concentrations below 3 pct, the effect of manganese is small. The values for activity coefficient of sulphur given in Table I were calculated from the following relations: S (in liquid metal) + H2 (gas) = H2S (gas) [l] K ph2s/?s X %S X phg = 0.00251 [2] where K is the equilibrium constant for the reaction, PH2S and ph2 are the partial pressures of hydrogen sulphide and hydrogen, respectively, and ?s is the activity coefficient of sulphur. The standard state for sulphur was taken to be a 1 pct solution of sulphur in pure iron. The numerical value for K at 1600°C was determined in the earlier work. For the purpose of showing graphically the results of the tests run at 1600°C, the activity coefficients of sulphur were recalculated so as to correspond to a manganese concentration in the metal of 2 pct in each case. In the calculation it was assumed that the increase in sulphur content of the metal at equilibrium caused by the presence of manganese

Jan 1, 1953

By H. A. Kuhn

THE Pittsburg coal field in western Pennsylvania, is conceded to be the most important in the world. To measure its importance it is necessary to understand the extent of its service in the various industries of the country. Probably 90 per cent. of the pig iron manufactured in the United States up to the present time has been made by using coke manufactured from the Pittsburg coal seam in western Pennsylvania. This coal field is the foundation on which the city of Pittsburg rests and is the reason for the great growth of the iron industry in the Pittsburg district. Iron ore is brought to this district, not because Pittsburgh is a. natural location, over other locations, for the iron and steel industry-as its position places the manufacturer at the mercy of the railroads-but the ore is brought 1,100 miles to meet the fuel. It can be said that the illuminating-gas industry in the United States has used this coal exclusively to the same extent that the pig-iron maker has used it. It may be also said that 20 to 25 per cent. of the fuel used on railroads in the United States conies from this coal field. The Pittsburg coal field is unquestionably the center of the industrial population of the United States, for in addition to the industries of the district and those closely adjoining, it has tributary to it all the cities and industries along the Great Lakes and practically all of Canada with the exception of the extreme western and eastern ends. It supplies the industries and population west of Duluth and Superior many hundreds of miles. This coal is floated down the Ohio and Mississippi rivers, supplying the towns en route, and is delivered in New Orleans, a distance of 2,200 miles, for approximately SO to 90 c. per ton transportation cost. It is delivered on the docks of Superior and Duluth at a cost of transportation 50 c. a ton less than the cost of transporting the same coal from Pittsburg to a local consumer in Philadelphia. With other Appalachian coals it has large markets east and along the seaboard, especially for byproduct-coke making and the illuminating-gas industry. It is considered the premier railroad fuel of the world on account of the fact that this coal in a given-sized locomotive will probably haul more cars than any other coal in the

Jan 10, 1914