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By W. N. Page

The Red-Ash colliery was the scene, March 6, 1900, of an explosion by which some fifty men were killed or wounded, and much property was injured or destroyed. On March 19th, the earliest date at which entrance was practicable, I examined the mine with the view of determining the extent, cause and point of origin of the explosion, and with the permission of Mr. F. Howald, the manager, and the owners of the property, for whom the examination was made, I present in this paper some results of that examination, comprising the essential portions of my report to them. Location and Plan of the Mine. The Red-Ash colliery, situated on the south side of New river, next below the Rush Run mine, with which the underground workings are connected, is a drift-mine, the entrance to which is on a slope, about 500 ft. above the river. (Fig. 1 is a plan of the workings, in which the points where human bodies were found are designated by numbers.)

Jan 1, 1901

By J. C. Bayles

An obstacle to the more careful and satisfactory study of metals has been the difficulty in harmonizing the results of chemical and physical tests. These give us records of observations made from different and often widely separate points of view, and while it is true that they furnish bases for more or less accurate triangulations into the realm of speculation, it is doubtful if, with only the laboratory and the testing machine, our investigations would not move in parallel lines, leaving between them an unexplored field in which must probably be sought the information which shall connect the results of chemical and physical tests, and give to both a practical value which neither has yet been found to possess. The use of the microscope in the study of metals is not a new thing, but it is only lately that it has begun to attract the attention it merits or to show results of tangible value to the metallurgist. Among recent valuable contributions to the literature of this subject, I have been especially interested in the work of Mr. A. Martens, of

Jan 1, 1883

By David L. Sikarskie

The measurement of residual stress fields using X-ray diffraction techniques requires the removal of successive layers of material from the specimen. In the process of removing these layers, the underlying residual stress field is altered; hence, if the original stress field is desired, corrections must be made to the measured stresses. Although these corrections already exist for a number of cases, an alternate procedure is outlined herein whereby the corrections are presented in series form. This form is numerically more convenient within certain ranges of correction depths and in its simplicity gives additional insight into the factors affecting the corrections. It, further, readily permits determining the proper slice depth to remove and the effect on the correction of both specimen thickness and magnitude and form of the measured residual stress itself. ONE of the techniques currently being used to determine residual stress fields is X-ray diffraction. Since X-rays are diffracted essentially from the layers of atoms near the surface of the specimen, of the order of 0.0001 to 0.001 in. depending on conditions, the method can only determine stresses* in this very thin surface region. Therefore, if the residual stress field at some depth into the specimen is required it is necessary to remove successive layers of material taking measurements after each cut. In the process of removing a thin surface layer of stressed material, however, the over-all equilibrium of a cross section of the specimen is disturbed with the result that the internal residual stress which existed before the slice of material was removed is altered. This alteration of stress occurs after each slice and its effect is cumulative. Thus, if the original stress distribution is desired, corrections must be applied to the measured stresses. These corrections can be found by satisfying equilibrium on each new cross section. These corrections have been derived in detail for a number of cases, e.g., cylinders, bars, plates,' and also are given in Appendix VI of Ref. 2. In each of the above cases the correction terms require the evaluation of definite integrals of the measured residual stress distribution. In general, this is indeed the way one would proceed with the corrections. There are a number of instances, however, where the integrals can be replaced by a series form of correction which is considerably more convenient numerically and, also, gives added insight into the factors which influence the corrections. Such a form of the correction is particularly useful, for example, if one is interested in designing a test such that stress corrections will be small. Also, the question of what depth of slice to use in relation to both specimen thickness and residual stress distribution itself must invariably be answered. Questions of this type can be readily answered using a series form of correction. In addition, if the correct residual stress distribution is needed only for limited depths into the specimen the series correction provides numerical results which are equivalent to the integral correction. I) DERIVATION OF THE SERIES CORRECTION The problem to be considered here is that of a bar or plate,* i.e., either one or two dimensions of the specimen is large compared with the third. Also, the residual stress distributions considered depend only on z, see Fig. 1. In view of this it can be shown' that a stress correction, based on simple beam theory, satisfies all conditions of "elasticity theory" and hence is the correct (and only) solution. The integral form of the correction (this assumes that the measured residual stress is a piecewise continuous function) as given in Ref. 1 is:

Jan 1, 1968

By John V. Beall

When a mine has ore averaging 5% lead, 7% zinc, 0.60% copper, 1/4 oz gold, and 6 oz of silver, adequate reserves, power and water, easy access to market, and is situated in beautiful natural surroundings with the advantages of a big city within a few hours drive, there is little left to be desired. Such a mine is the Mayflower mine of the New Park Mining Co. situated 30 miles from Salt Lake City in the Wasatch Mountains near Keetley, Utah.

Jan 1, 1949

By N. A. Gokcen

The construction and operation of a simple and accurate vacuum-fusion apparatus are described in detail. Absolute accuracy of the oxygen analysis has been determined by the reduction of oxides weighed to 20.01 mg. The Bureau of Standards steel samples have been analyzed repeatedly, and their inadequacy for the acceptable range of oxygen is presented. The limits of interference of manganese, aluminum, and titanium, and the effects of tin and water-cooling of the furnace tube have been investigated in detail. ANALYSIS by the vacuum-fusion method consists of melting a sample in a degassed graphite crucible under high vacuum and extracting oxygen as carbon monoxide, and nitrogen and hydrogen as gaseous elements. The resulting gas mixture is analyzed for its constituents, from which oxygen, nitrogen, and frequently hydrogen may be determined within 1/2 hr or less. The history and summary of methods1-4 and extensive reviews of literature'.' for the determination of gases in metals have been published in detail. The foundation of the modern vacuum-fusion method of analysis was laid by Jordan and Eckman,11 and Oberhoffer and his associates.1, 10 Diergarten,11, 12 Meyer,13 Thanheiser et al.,14, 15 Ericson and Benedicks,16 Ziegler,17, 18 Sloman,16-23 Thompson et al.,24 and later many others contributed to the improvement, accuracy, and limitations of the method. In some investigations, however, the limitations of the procedure and the errors involved in the analysis require further critical examination and evaluation. In many apparatuses, lunnecessarily complicated and cumbersome features and elaborate precautionary measures do not have conclusive advantages. The purpose of this invest:igation was, therefore, a) to construct a very simple and accurate apparatus, b) to determine the absolute accuracy of results, c) to establish the limits of interference of manganese. aluminum, and titanium, d) to reexamine critically the reduction and recovery of oxygen from oxide powders of various sizes, and e) to evaluate critically the eight steels of the Bureau of Standards."' Apparatus The apparatus used in this investigation is shown in Fig. 1. It has been used for over three years, during which several modifications were made. A brief description of the apparatus is as follows. The transparent silica or Vycor tube, E, is joined to the Pyrex head, B, by mean of a face to face ground joint sealed with mercury. A similarly sealed ball and socket joint would also have been satisfactory. A standard tapered ground joint, successfully used by some investigators, was tried, but in the absence of grease or vacuum cement it was very difficult to disassemble the joint. The graphite crucible, G, 15/16 or l 1/8 in. OD, 2 1/2 or 3 in. high, 1/10 in. wall, shown enlarged in Fig. 3a, is made from the rods containing less than 0.08 pet ash. Two baffles, overlapping sufficiently to prevent spattering, are inserted in two opposite slots cut on a band saw. Two small holes in the baffles permit the temperature measurement of the melt. A second type of crucible, Fig. 2, G, shown enlarged in Fig. 3b, is also satisfactory, though not as convenient as the crucible shown in Fig. 3a. Somewhat similar but more elaborate crucibles with baffles were first tried by Ericson and Benedicks,'" but were abandoned in favor of a crucible with a stopper and a peripheral graphite powder filter. Various versions of their crucible with a stopper have later been used by other analysts, 10-27, 27-30 The crucible is packed directly in E, Figs. 1 and 2, with —35 + 48 graphite powder for shielding and insulation. A minimum layer of 7 mm of powder is necessary for keeping E cool and minimizing the blank. The first use of graphite powder was made by Sloman,10-23 Who found that —200 mesh was the most satisfactory. The author tried various size powders from +20 to —100 mesh and found that a) particles of 28 mesh or larger were not sufficiently insulating and were heated by the induction current, b) the powder finer than 60 mesh packed too much and did not readily permit the escape of gases even when the crucible was heated very slowly; hence, the powder was occasionally blown out of the side in the crucible and in the mercury pump. The optimum size powder is, therefore, 35 to 48 mesh size or 0.4 mm in average diameter. The furnace is designed to eliminate excessive amounts of

Jan 1, 1959

By F. W. Warnaars

The Los Bronces copper deposit is located on the west side of the Andes Mountains in central Chile about 54 km (34 miles) northeast of Santiago. The deposit consists of a hydrothermal breccia complex that was emplaced in the western part of an earlier porphyry copper mineral system. This system exhibits propylitic, sericitic, silicic to potassic hydrothermal alteration with disseminated and stockwork copper-molybdenum sulfide mineralization over an area of about 12 km2 (5 sq. miles). The Rio Blanco Mine, about 800 m (0.5 miles) east of Los Bronces, is an underground block caving mine located in the central part of this system. The mineralization at Rio Blanco is characterized by one of the highest primary copper grades known for porphyry copper systems. Present mining at Los Bronces is from an open pit located in the northern part of a hydrothermal complex that is composed of at least seven different copper-bearing tourmaline breccia pipes. Together these pipe form one large contiguous body about 2 km (1.2 miles) long and 0.7 km (0.4 miles) wide at the surface. This body crops out at elevations of between 3.5 and 4.1 km (2 and 2.5 miles) above sea level. Each breccia pipe has a characteristic shape, matrix, type and degree of mineralization and alteration. Quartz, tourmaline, specularite, anhydrite, pyrite, chalcopyrite, sericite, chlorite, and rock flour are present in significant quantities in most breccia matrices. The clasts consist mainly of quartz monzonite, andesite, and minor amounts of quartz latite porphyry. The seven different breccia pipes are identified, from oldest to youngest, as Ghost, Central, Western, Infiernillo, Fine Gray, Anhydrite, and Donoso. The breccias coalesce, interfinger, or display gradational contacts. In places they form sharp contacts with each other and with the surrounding intrusive or volcanic host rocks. Early breccias are disrupted or cut by subsequent breccia events. Explosive emplacement was followed by collapse after the release of hydrothermal pressures. The Los Bronces breccia complex has sharp contacts with the surrounding intrusive stock and andesites. These contacts dip inward on the north, west, and south margins. The eastern contact is nearly vertical and/or dips slightly outward and suggests a westward tilting of the body of about 15° after emplacement. The breccia complex is not fully explored at depth, but is known to extend at least below the 3 km (2 mile) elevation in its central eastern part, about 1 km (0.6 miles) below the highest breccia outcrop. Primary copper sulfide minerals are most abundant in the Donoso and Infiernillo breccias where the matrix consists mostly of tourmaline, quartz, sulfides, specularite, and/or chlorite with minor amounts of rock flour, sericite, and clay. The volume of the matrix material is mostly smaller than in the other breccias. Secondary enrichment enhanced the primary copper grade in the southern two-thirds of the Los Bronces breccia complex and in much of the surrounding porphyry system. Primary grades of 0.3% to 0.5% copper have been enriched to approximately 0.9%. This enrichment may extend from surface outcrops to a depth of more than 500 m (1, 640 ft) in more permeable parts of the breccia complex. A leached and oxidized capping of variable thickness covers a large part of the enriched layer and exceeds 300 m (984 ft) at one location. The shape and depth of the enrichment blanket and the capping suggest that the enrichment process is still active.

Jan 1, 1983

By Oscar M. Wicken

The mineral magnesite (MgCO3) if pure would consist of 47.7 pct MgO and 52.3 pct CO2. It is one of the calcite group of rhombohedral carbonates which includes calcite (CaCO3), siderite (FeCO3), rhodocrosite (MnCO3), and smithsonite (ZnCO3) among others. The members of this group enter into a wide range of substitutional solid solutions, whose range is still imperfectly known.* MgCO3 and FeCO3 appear to form a complete series of which breunnerite (ferroan magnesite) is a well known variety. Because of the considerable difference in size of the Ca and Mg ions, there is only a limited range in substitution in the MgCO3-CaCO3 series. (Dolomite, not strictly a member of the calcite group, may be regarded as an ordered structure developed when Ca: Mg is approximately 1:1.) The practical consequence of these relationships is that the lime carbonates (calcite and dolomite) are commonly present in magnesite deposits chiefly as mechanical mixtures. Magnesite may be either crystalline or "amorphous" (cryptocrystalline). The crystalline form has a hardness of 3.5 to 4.0. The color may range from white to black through shades of yellow, blue, red or gray. It is rarely found pure but generally contains variable amounts of iron, lime, and silicate minerals. Color cannot be used as an index of purity except under specific local conditions in a single quarry. In a given quarry an experienced man may be able roughly to grade magnesite by intuitively assessing such features as crystallinity and color. Crystalline magnesite occurs in larger deposits, but the cryptocrystalline variety is more common. The latter is fine-grained and compact and shows no cleavage; often it is descriptively called "bone magnesite." The fracture is usually conchoidal, and hardness is 3.5 to 5.0. It is normally white, but sometimes it is light yellow to orange or buff because of included impurities. In addition to lime and iron, which may be present in small amounts, silica may be present as inclusions of serpentine, quartz, chalcedony, or other minerals. In general, commercial deposits of "amorphous" magnesite have lesser amounts of accessory minerals than does crystalline magnesite. The specific gravity of "amorphous" magnesite is 2.90 to 3.0. Pure, crystalline magnesite has a specific gravity of 3.02, but iron carbonate, commonly present, raises the specific gravity to somewhat higher values. Magnesite dissociates, when sufficiently heated, into magnesia (MgO) and carbon dioxide. Either simultaneously or upon further heating the magnesia develops a crystalline structure identical with that of the natural mineral, periclase, which occurs sparingly in nature but not normally in workable deposits. Brucite is another magnesium mineral capable of being directly reduced to magnesia by heat alone. Brucite (Mg(OH)2) has been mined for the production of various grades of magnesia. It theoretically contains 69.1 pct MgO and 30.9 pct H2O. It may be white, but is generally blue or green with a gray cast. It has a translucent appearance. The mineral is soft (hardness, 2.5), and has a specific gravity of about 2.4. Nomenclature of Products The word "magnesite" literally refers to the natural mineral, but common usage has applied it plus a prefatory word to two other

Jan 1, 1960

By John Chipman, Minu N. Dastur

This paper presents equilibrium data on the reaction of water vapor with vanadium dissolved in liquid iron at 1600°C. The thermo-dynamic behavior of vanadium and oxygen when present together in the melt is discussed. A deoxidation diagram is presented which shows the concentrations and activities of vanadium and oxygen in equilibrium with V209 or FeV2O4. STUDIES of the chemical behavior of oxygen dissolved in pure liquid iron1-3 have served to determine with a fair degree of accuracy the thermody-namic properties of this binary solution. The practical problems of steelmaking, however, involve not the simple binary but ternary and more complex solutions. Only a beginning has been made toward understanding the behavior of such systems. The silicon-manganese-oxygen relationship was studied long ago by Korber and Oelsen4 and more recently by Hilty and Crafts." The carbon-oxygen reaction was investigated by Vacher and Hamilton6 and by Marshall and Chipman.7 A number of deoxidizing reactions have been studied empirica1lys'10 with the object of determining the appropriate "deoxidation constants." The work of Chen and Chipman" afforded a clear-cut view of the effect of the alloy element, chromium, on the thermodynamic activity of oxygen in liquid ternary solutions. These investigators determined the oxygen content of experimental melts which had been brought into equilibrium with a controlled atmosphere of hydrogen and water vapor and were able to show that the presence of chromium decreases the activity coefficient of oxygen. They determined also the conditions under which the two deoxidation products, Cr2O3 and FeCr2O4, were formed and showed that the activity of residual oxygen is considerably less than its percentage. It was the object of this investigation to apply a similar method to the study of molten alloys of iron, vanadium, and oxygen. Vanadium was once considered a moderately potent deoxidizer, but this is now known to be erroneous, in the light of its behavior in steelmaking practice. Its reaction with oxygen retains a certain amount of practical interest in that a high percentage of one element places a limit on the amount of the other that can be retained. As a deoxidizer it will be shown that vanadium lies between chromium and silicon. Experimental Method The apparatus was that used by the authors3 in their study of the equilibrium in the reaction: H2(g) +O = H2O(g);K,= [1] PII., ao Crucibles of Norton alundum or of pure alumina were used. The latter were made in this laboratory and were of high strength and low porosity. Under conditions of use they imparted no significant amount of aluminum (less than 0.01 pct) to the bath. Temperature measurements were made with the optical equipment and calibration chart of Dastur and Gokcen.= The charge was made up of calculated amounts of ferrovanadium (20 pct V) and clean electrolytic iron totaling approximately 70 g. The first few heats were made in alumina crucibles with an insufficient amount of vanadium so that no oxide of vanadium would be precipitated under the particular gas composition. All the heats were made at 1600 °C under a high preheat and with four parts of argon to one part of hydrogen in the gas mixture to prevent thermal diffusion. The rate of gas flow was maintained constant at 250 to 300 ml per min of hydrogen. The time for each heat was three quarters of an hour after the melt had melted and attained the required temperature (1600°C). The water-vapor content of the entrant gas mixture was gradually raised in succeeding heats, keeping the vanadium content of the melt constant. This was controlled by manipulation of saturator temperature. A point was reached when for a given H2O:H2 ratio some of the dissolved vanadium was oxidized and appeared as a thin, bright oxide film on top of the melt. By raising the temperature of the melt it was possible to dissolve the oxide film which reappeared as soon as it was cooled down to 1600°C. The temperature readings taken on the oxide film were consistently higher by 80" to 85 °C as observed by the optical pyrometer. The heat was allowed to come to equilibrium under a partial covering of this oxide film. At the end of the run the power and preheater were shut off and the crucible containing the melt was lowered down into the cooler region in the furnace. This method of quenching proved quite

Jan 1, 1952

By A. H. Daane, A. S. Wilson

The U-Cr system is of the simple eutectic type with some solid solubility of chromium in r and ß uranium. The eutectic occurs at 20 atomic pet Cr and melts at 859°C. The maximum solubility of chromium in y uranium is 4 atomic pet at the eutectic temperature, and in ß uranium the solubility is estimated to be 1 atomic pet. y-ß and ß-a transformations were found to occur at 737°estimatedt° and 612°C respectively. DURING 1944 and 1945, the U-Cr constitution diagram was studied in this laboratory as a part of a research program on uranium metallurgy in the Manhattan Project, and the work was described in a Manhattan Project report issued in December 1945. This paper is based on that report, which has been declassified. Prior to this study, it had been shown by other Manhattan Project workers that the low Cr-U alloys could be quenched to retain the form of uranium. Experimental The uranium used in this work was massive metal prepared in this laboratory and contained less than 0.1 pct of other elements. The chromium was 200 mesh powder obtained from the A. D. McKay Co. and was found on analysis to be 99.5 pct Cr with 0.3 pct Fe the major impurity. Alloys, weighing 400 to 600 g, were prepared by induction heating the components to 1700°C in slip-cast ZrO, crucibles in a vacuum of 3x103 mm Hg. TO prevent too violent agitation of the melt by the induction field with subsequent crucible breakage and sample loss, the ZrO2 crucible was placed in a graphite crucible, which was surrounded by a layer of powdered carbon insulation 2 to 3 cm thick. Polished vertical sections of the alloys were examined microscopically to confirm their homogeneity. Heating and cooling curves were taken on the alloys by reheating them in ZrO, crucibles to 1200°C and inserting a mullite-protected chromel-alumel thermocouple into the melt by means of a slip seal in the vacuum head of the furnace. A recording potentiometer traced the curves which had a slope of from 3" to 6" per min. Samples of the alloys were prepared for metallo-graphic examination by conventional mechanical polishing techniques followed by an electrolytic polish in an ethylene glycol-phosphoric acid-ethyl alcohol bath. The structure of the alloys was brought out clearly by this procedure so that no further etching was required. Samples for chemical analysis were taken from drillings from the top, center, and bottom sections of the alloys. The uranium was determined by titra-tion with Ce(SO1)2, while the chromium was titrated with FeSO,; the uranium and chromium totaled at least 99.6 pct in all of the alloys prepared. X-ray samples were prepared by filing bulk specimens in a helium-filled glove box and annealing the resulting powder in a zirconium-gettered helium atmosphere. A 114.6 mm diam Debye-Scherrer camera and a Weyland nonsymmetrical self-focusing camera were used with filtered copper radiation to obtain the powder X-ray diffraction data. Results The data obtained in this study have been combined to construct the constitution diagram of the U-Cr system shown in Fig. 1 where the arrests observed in cooling curves are indicated by dots. The liquidus arrest was quite distinct in thermal data taken on alloys in the range 0 to 20 atomic pct Cr. The eutectic arrest was not observed in studies on the 2.5 and 4.5 pct Cr samples but appeared in the 7.5 pct samples, which suggested some solubility of chromium in y uranium. On quenching from 859 °C, the 2.5 pct sample showed but one phase while the 4.5 pct sample contained a small amount of the eutectic along the grain boundaries; see Figs. 2 and 3. From this the maximum solubility of chromium in r uranium has been set at 4 pct. X-ray studies on these samples showed that the r phase was not retained at room temperature by quenching, but in each case a pattern was observed .which has been identified with the ß phase of uranium. Thermal data show the y-ß transformation of uranium lowered to 737°C as a consequence of this solubility. On quenching from the ß range (660°C), precipitation of chromium in the primary uranium is observed in the 2.5 and 4.5 pct Cr samples (see Figs. 4 and 5),

Jan 1, 1956

By R. E. Steiner, A. F. Weinberg, R. J. Van Thyne

Uranium solubility in molten bismuth was determined in the temperature range 350" to 600°C, varying from 0.09 wt pct to near 2 wl pct, respectively. Zirconium and magnesium, simulated fission products, and sodium were found to increase or decrease uranium solubility in bismuth in the temperature range 370° to 420°C, depending upon the addition or temperature. Data were obtained by both chemical analysis of samples obtained by filtration from saturated melts and by measurement of the variation of electrical resistance with temperature. IN a study for the United States Atomic Energy Commission, a Liquid Metal Fuel Reactor Experiment (LMFRE) was proposed in which a solution of uranium dissolved in molten bismuth served as both the fuel solution and the primary coolant. It was proven necessary to add other elements to this solution for certain desirable properties: 1) zirconium acts to prevent reaction of the fuel solution with the graphite core (which would result in the formation of uranium carbide) and also to inhibit mass transfer of iron and chromium from the walls of the reactor tubing to cold spots resulting in blockage; and 2) magnesium acts as a deoxidant and prevents loss of uranium. In addition to these intentional additives, other elements entering the solution are: 1) fission products which are a by-productof reactor operation, 2) sodium (the secondary coolant) through possible small leaks in the heat exchange system, and 3) iron and chromium from the container walls. Each of these elements may affect the solubility of uranium in the fuel solution. Since the concept of this reactor is based upon a liquid solution, the relatively low solubility of uranium in bismuth will dictate the minimum temperature at which the reactor heat exchanger can be operated. The solubility of uranium in liquid bismuth and the effect of some additives on this solubility have pre- viously been investigated1-' (more recent results6" became avaliable after initiation of this investigation). However, some discrepancies between the various results have been noted, and also the cumulative effect of the several additions or impurities was not known. This investigation was commenced to obtain definitive results. MATERIALS Analyses of the materials used are presented in Table I. Although ostensibly of extreme purity (99.999 wt pct), the bismuth was found to contain a significant amount of impurities, probably oxides. Except for determination of the binary system, Bi-U, prior to use all bismuth was hydrogen-refined and

Jan 1, 1962

The early history of coal in West Virginia is all included with that of Virginia in the few records available but for present-day readers it is much more convenient that the account of this area prior to the Civil War be grouped with that since then. West Virginia shares with Maryland in the first record of coal noted in the Appalachian fields. This was along the upper Potomac River, in the fall of 1736, and the data about it are given under Maryland, as most' of the development of the area was in that state. The next record of coal was left by John Peter Salley in his account of the expedition he made with Howard and others in 1742. (See p. 17.) The next reference found is that of Dr. Thomas Walker, in 1750, when he passed through the Pocahontas coal field. This has been described on page 18. Six years later, in March 1756, Capt. William Preston, who was with Major Lewis in his expedition down Big Sandy River, wrote in his journal: "Tuesday 2.... Moved down the branch and- came to the main creek where we camped. Put on half rations. Came into the Cole land; crossed the river 8 times."1 The party crossed from the head of Clinch River to Horse Pen Creek, then to Jacob's Fork and down it to Dry Fork of Tug River; the coal was seen in what is now McDowell County. The early settlers must have known of the presence of the great Pittsburgh bed in the northern part of the state, both along the Monongahela River and elsewhere, because the outcrops could easily be seen along streams in many. places. The report of the Commissioners for Adjusting Claims to Unpatented Lands, made in 1789, contains several references to coal in the names of streams, made years earlier. The earliest is one for 1,000 acres to James Walker, in Monongalia County, "adjoining his settlement on a branch of the right hand fork of Pappaw known by the name of the Stone Cole Lick made in the year 1773."2 Another of 400 acres in the same locality was settled upon in 1777.2 A third spoke of a tract "on the left hand fork of Stone Cole about three miles from the mouth," in 1783.1 Monongalia County then covered all of the northern

Jan 1, 1942

By Kenneth A. Gutschick, Robert S. Boynton

Lime has become a very general loosely used term that includes virtually all finely divided types of limestone as well as burned forms of lime. However, actually (and according to Webster) lime is only calcined limestone (known as quicklime, calcium oxide, or calcia). It also embraces the secondary product of quicklime-namely hydrated lime or slaked lime (calcium hydroxide). Fundamentally there are two basic types of limestone-high calcium and high magnesium (dolomitic). The dolomitic quicklimes and hydrated limes correspond to their "high calcium" counterparts in terminology. The only difference is that the dolomitic types are a combination of the elements, calcium and magnesium, whereas high calcium limes contain only 0.5 to 2 pct of magnesium. Both types, which are made through calcination at high temperatures, are shown chemically as follows: [ ] In the above reactions either type of limestone is burned (calcined) in furnaces known as lime kilns so that the carbon dioxide content of the stone is expelled as a gas. With high calcium limestone 44 pct of the weight of the stone is lost; with dolomitic limestone the loss is 48 pct. Also, both reactions are chemically reversible since quicklime absorbs carbon dioxide readily and in so doing transforms itself back to the original carbonate form. Most quicklimes correspond in size to the original rock, which varies from one to eight inches, depending upon the type of kiln employed. Slight shrinkage occurs during calcination. A few limestones when calcined will decrepitate (break down) into a fine powdery form; however, chemically it is the same as the pebble or lump-sized material. For some uses the lumps of quicklime are pulverized and packaged. Since quicklime has great affinity for moisture and carbon dioxide, it is perishable. When it absorbs moisture and CO2 slowly from the atmosphere, the resulting product is "airslaked." Consequently, it must be stored in dry, moisture-proof areas that are free from CO2. Because of its fugitive nature, it is usually consumed within a few weeks after manufacture. A more stable form of lime is hydrated lime-commercially made by adding water to the quicklime until the lumps break down into a fine, white powder. This product, which is usually packaged, can be stored more safely; and many of its uses are interchangeable with quicklime. The irreversible chemical reactions of hydrated lime for both high calcium and dolomitic types are: [ ]

Jan 1, 1960

IN A broad sense, the subject of slag control includes not only the adjustment of the composition of the slag but also of its relative weight in terms of percentage of the metal-bath weight. The slag is made up essentially of (1) oxides formed by oxidation of silicon, manganese, phosphorus, and iron in the charge, (2) SiO2, MnO, and other oxides charged as such, and (3) lime charged or fed to the heat to neutralize the “acidic”: oxides present (Si02 and P2O5). Obviously, if practically perfect control were possible over the composition of the charge as well as over the oxidation process during the melting period, both slag weight and composition would be very nearly predetermined, and the control problem would be solved. Unfortunately, we have only a somewhat. imperfect control over these factors, so that beginning at about the lime-boil period in the heat, and later in the early refining period, it becomes necessary for the first helper, as the heat progresses, to make quick estimates of the composition and physical state of development of the slag (as well as of the amounts of phosphorus, sulphur, and carbon in the metal) from which he can judge how much extra lime, roll scale, fluorspar, and other additions will be needed not only to give the desired final slag composition, but also to hasten the complete fusion or shaping to a desired viscosity in the liquid state. The discussion in this chapter is limited essentially to those more or less practical rapid methods of slag estimation that have been used, or at least tried, for these purposes under actual open-hearth shop conditions. Another aspect of importance here is the extent to which the freedom of choice to the operator is limited by available charge materials, product specifications, and the oxidation and other chemical- reaction effects which inevitably occur under the conditions of this furnace process. For a better understanding of these control limitations and of the chemical and physical background of the subject, the reader is urged to read at least part of the discussion in Chapters XVII and XIX. Charge materials are discussed in Chapter IV, and many of the details of charging, melting, and refining practice closely related to slag control are given in Chapters VI and VII.

Jan 1, 1944

By George A. Connell

BORAX, a sodium borate and the principal sodium salt of boric acid, has been surrounded with romance and with a certain amount of mystery. Its early history is not entirely known but it has been contended that Marco Polo in his travels in the Orient introduced this material to the West during the Middle Ages. Without doubt its first source was from Central or Eastern Asia and perhaps from Tibet. The well-known fluxing properties of borax probably introduced its earliest use that of a soldering flux in the manufacture of jewelry. This use, as well as the remote location of most borate deposits, added greatly to the romance associated with the thought of borax. COMPOSITION The common element in all borates is boron but this element has never been discovered as such in any deposit, because of its great affinity for oxygen. Incidentally, this makes its manufacture extremely difficult and expensive. Elemental boron undoubtedly has many properties that make it quite valuable but the extreme cost involved in producing it from its oxidized form prevents its general use or manufacture. The common element in all borates is boron but this element has never been discovered as such in any deposit, because of its great affinity for oxygen. Incidentally, this makes its manufacture extremely difficult and expensive. Elemental boron undoubtedly has many properties that make it quite valuable but the extreme cost involved in producing it from its oxidized form prevents its general use or manufacture. The common element in all borates is boron but this element has never been discovered as such in any deposit, because of its great affinity for oxygen. Incidentally, this makes its manufacture extremely difficult and expensive. Elemental boron undoubtedly has many properties that make it quite valuable but the extreme cost involved in producing it from its oxidized form prevents its general use or manufacture of disodium tetraborate (Na2B4O7.10H2O), and, further, can be considered as the acid salt of monosodium metaborate. However, the term "acid salt" might imply that borax gives an acid reaction in solution, which is not true, since solutions of borax are decidedly alkaline, giving a pH of 9.2. Boric acid, also known as boracic acid, has the chemical formula H3BO3; it occurs in nature as the mineral sassolite. Sassolite has been

Jan 1, 1949

Cortez Gold Mines operated a conventional 2,100 mtpd (2, 300 stpd) cyanidation plant until the 5 million ton ore body was worked out in 1973. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME 1975.) Sub-grade material was dump leached at the mine -mill location in 1971, and the resulting pregnant solutions were added to the feed in the plant's Merrill-Crowe precipitation unit for gold recovery. In early 1973, however, Cortez put into operation their Gold Acres mine which was located about 13 krn (8 miles) to the west of the 'Cortez mine. Economics at that time dictated the heap leaching of the lower grade portion of the ore body with the higher grade being trucked to the Cortez plant for milling. Leach ore was moved directly from the open pits to prepared leach pads located on a gently sloping terrain of 570 grade. Preliminary pad preparation at the leach site included removal of vegetation which was composed of sagebrush and grading of the base. Pads were comprised of mostly minus 74 micron (200 mesh) materials and were made of clay-silt or slime tailings spread on the pad area in thin 50 to 75 mm (2 to 3 in. ) layers, compacted with a vibrating roller compactor to a depth of about 380 mm (15 in.). This was followed with a 100 mrn (4 in.) layer of coarse rock to prevent erosion on the pad and to allow solutions to move laterally under the heaps. Leach solutions normally traveled the length of the heaps and were collected in plastic-lined ditches from which they flowed by gravity to the pregnant solution storage pond. Both pregnant and barren ponds were constructed in the same manner as the leach pads - being lined with compacted tailings or clay silts. The Gold Acres facility required pregnant solution storage of 6 million liters (1.6 million gal) and barren storage of 3.8 million liters (1.0 million gal) designed to handle two to three days of heap drainage or to retain run-off from a major rainfall. Various heap heights were tried including 3, 6, and 9 meters (10, 20, and 30 ft), with a 6 m (20 ft) height selected. Ore was dumped by trucks in a single 6 m (20 ft) lift using a D8 Caterpillar dozer for

Jan 1, 1981

By R. S. French, W. R. Hibbard

M. Balicki—As one who some years ago spent much time searching for an alloy with high thermoelectric power that would be suitable for heat energy-electric energy converter based on the principle of a thermoelectric pile, I would like to make a few comments touching upon that aspect of this excellent paper. The specifications which one of the authors has formulated so aptly do not appear to me as complete, and I would like to add to them the following two demands: (1) Melting point not less than 500°C and (2) elongation on 2 in. at least 5 pct. Condition 1 would insure good thermodynamic efficiency of the pile by permitting a reasonably high temperature of the hot junction and at the same time making it suitable for such application as salvage of vast amounts of energy which is lost in the form of sensible heat of waste gas leaving industrial and other furnaces. Condition 2 is also a very important one; particularly, because of the fact that all elements and alloys which exhibit attractive thermoelectric properties seem to be as a rule very brittle. The design, erection, and maintenance of a pile incorporating a brittle material in one of the most crucial parts of the converter would be a very difficult and costly affair indeed. To illustrate this point I may mention great difficulties I encountered when trying to make thermocouples of silicon. Casting of 1/4-in. rods in iron or graphite mould yielded only broken pieces which had to be welded into a bar. Such a bar had to be handled very carefully during the determinations. Raising the molten silicon by vacuum into a silica tube and solidifying it there solved the problem of stability as well as that of electrical insulation of a bar. On tests this bar (98.5 pct Si) showed dE/dT = — 275.5 microvolts per degree C vs. copper. Thermocouples made of this grade of silicon and pure nickel developed 0.39 v when temperature of the hot junction was approximately 1000°C. The possibility Dr. Gonser has mentioned of trying to impart to metals good thermoelectric properties by alloying them with silicon, has been considered. The preliminary survey which I have made was, however, disappointing. Perhaps, it will be advisable to record the results obtained (table VI), as this might save repetition of search in fields already covered.

Jan 1, 1951

By R. W. Smith, G. W. Delamore, J. A. Spittle

RECENT work by the authors1 has demonstrated that the crystals constituting the equiaxed zone in small ingots originate, on quenching, in the chilled liquid. Whether any partial remelting2 takes place of the den-drites nucleated in the chilled region was not examined. However, the claim3 that this normally occurs, leading to crystal multiplication, has been disputed.4 Thus, for an equiaxed zone to form in the central region of an otherwise columnar ingot, it is necessary to transport crystallites from the chill zone to this region. In view of the work of Uhlmann et al.,5 it was considered desirable to examine the transport process further. Experimental Technique and Results. In the earlier experiments by the authors.1 a stainless-steel crucible 17/8 in. diam, 41/2 in. high, ad containing an alloy of A1-4 wt pct Cu, was quenched from 660°C into a bath of Wood's metal maintained at 100°C. Under these conditions a central zone of equiaxed crystals was obtained. However, if a 2250-G unidirectional magnetic field was applied for 2 sec immediately following the instant of quenching, or subsequently, the equiaxed zone did not appear. Careful thermal analysis showed that the short application of the field had virtually no effect on the thermal conditions of the melt. It thus appeared that the crystallites which were ultimately to grow large enough to form the equiaxed zone existed in the liquid immediately ahead of the interface, occlusion being prevented by the convective fluid flow induced by freezing from the walls of the cylinder. When this relative movement was prevented by the Lorenz forces induced by the application of the magnetic field, the crystallites were caught up in the rapidly advancing dendrite solid-liquid interface. To test this assumption, a small ingot of super-purity aluminum was partially melted and poured into a heated stainless-steel mold which was then immediately quenched into molten Wood's alloy. The resulting structure is shown in Fig. 1, and consists of columnar grains with a central fine-grained zone. Thus the unmelted solid present at the beginning of solidification had been transported ahead of the columnar interface. Discussion. There appears to be little or no information available on the interaction of suspended particles with a metallic solid/liquid interface. However, some information is available for transparent organic sys-

Jan 1, 1969

By B. D. Cullity, J. T. Norton, M. Telkes

M. Balicki—As one who some years ago spent much time searching for an alloy with high thermoelectric power that would be suitable for heat energy-electric energy converter based on the principle of a thermoelectric pile, I would like to make a few comments touching upon that aspect of this excellent paper. The specifications which one of the authors has formulated so aptly do not appear to me as complete, and I would like to add to them the following two demands: (1) Melting point not less than 500°C and (2) elongation on 2 in. at least 5 pct. Condition 1 would insure good thermodynamic efficiency of the pile by permitting a reasonably high temperature of the hot junction and at the same time making it suitable for such application as salvage of vast amounts of energy which is lost in the form of sensible heat of waste gas leaving industrial and other furnaces. Condition 2 is also a very important one; particularly, because of the fact that all elements and alloys which exhibit attractive thermoelectric properties seem to be as a rule very brittle. The design, erection, and maintenance of a pile incorporating a brittle material in one of the most crucial parts of the converter would be a very difficult and costly affair indeed. To illustrate this point I may mention great difficulties I encountered when trying to make thermocouples of silicon. Casting of 1/4-in. rods in iron or graphite mould yielded only broken pieces which had to be welded into a bar. Such a bar had to be handled very carefully during the determinations. Raising the molten silicon by vacuum into a silica tube and solidifying it there solved the problem of stability as well as that of electrical insulation of a bar. On tests this bar (98.5 pct Si) showed dE/dT = — 275.5 microvolts per degree C vs. copper. Thermocouples made of this grade of silicon and pure nickel developed 0.39 v when temperature of the hot junction was approximately 1000°C. The possibility Dr. Gonser has mentioned of trying to impart to metals good thermoelectric properties by alloying them with silicon, has been considered. The preliminary survey which I have made was, however, disappointing. Perhaps, it will be advisable to record the results obtained (table VI), as this might save repetition of search in fields already covered.

Jan 1, 1951

By J. H. Hamilton, J. R. Lewis, J. C. Nixon, C. L. Graverson

Recently there has been much speculation concerning the advantages of using oxygen enriched air or pure oxygen in pyrometallurgical processes. The advantage of using oxygen in the iron blast furnace and elsewhere in ferrous production metallurgy has been given careful consideration and, as a result, several experiments have been tried and a number of papers have recently appeared on this subject.1,2 The use of oxygen in nonferrous smelting likewise poses interesting problems, therefore any experimental data obtained in this area should likewise be of general interest.3 Ashcroft4 in 1933 claimed to have obtained improved results when copper concentrates were roasted in pure oxygen. However, his paper gives neither data nor experimental details. A few years ago a research program was started in this laboratory which had for its objective the determination of the mechanism of oxidation of metallic sulphides under varying conditions. To date we have studied the oxidation of galena,5 sphalerite, and chalcocite. It is a well known fact that copper sul-phde minerals are oxidized slowly by weathering agents, so that the sulphides of copper which find their way into a tailing pond are gradually changed into water-soluble and acid-soluble compounds. Commercially, the oxidation of copper sulphide minerals is accomplished by roasting them at elevated temperatures in air using a suitable (usually multiple hearth) roaster. Most copper smelters are equipped with such roasters. In this phase of the study we were especially interested in determining the effect of temperature on: (I) the air oxidation of chalcocite and (2) the oxidation of chalcocite using commercial oxygen. The amount of sulphur removed as sulphur dioxide, the percentages of water-soluble copper, and of acid-soluble copper were determined for the air oxidation tests. In the tests using commercial oxygen, the percentages of sulphur dioxide, water-soluble copper, acid-soluble copper, water-soluble sulphur and acid-soluble sulphur were determined. For these experiments the temperature was varied from 250 to 700°C. Experimental Preparalion of Chalcocite (Cu2S) Natural chalcocite is not pure, Minerals such as bornite, covellite, and chalcopy-rite may be present. It was decided, therefore, to prepare some chalcocite in our own laboratory. An intimate mixture of sulphur and hool divided metallic copper in the ratio of 4 g of copper to I g of sulphur was prepared. TO this was added a slight excess of

Jan 1, 1949

By J. F. Keefe, C. J. Bier, O&apos

This study was carried out with copper because it represents the simplest form of sintering, in that but one metal is involved and all reactions occur in the solid state. The present work will cover three variables affecting the sintering process; namely, briquetting pressure, sintering temperature and time in hot zone. Although strength and dimensional changes were observed, this is primarily a metal-lographic study of the sintering process in copper. The term "sintering" should be understood to mean the process by which the metal powder particles of the compact become united into one body. Material Used The copper powder used was made from reduced copper oxide, and is typical of a grade used for the mass production of parts formed of metal powders. The screen analysis of the powder was as follows: Per Cent Per Cent BY BY Particle Size "Weight Particle Size "Weight Mesh Microns /80.... 0.00 43/32.... 17.49 80/I00..... 32/20......16.68 100/140.... 0.34 20/10.... 23.14 140/200___ 15.25 10/5..... 4.64 200/230....... 5.00 5/0...... 0.75 230/325.... 16.70 — 325....62. 69 The minus 325-mesh analysis was made by using a Roller Air Analyzer.1,2 The apparent density is here defined as the weight per cubic centimeter of the loose, unpacked powder as measured with the Hall flowmeter. The value for the powder used was 2.57 grams per cubic centimeter.

Jan 1, 1945