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By R. J. Grace

Primarily, this report presents the results of direct efficiency tests which were conducted on the 200-hp Coal-Pak, hot water generator, No. 3 unit, at the Otterbein College heating plant in Westerville, Ohio. Also included is a comparison of fuel consumption of the old plant with that of the new plant, a comparison of efficiencies obtained by the indirect method on a 300-hp Coal-Pak, and the cost of gas vs coal to show the savings effected by the latter. Tests were conducted with ten different bituminous coals selected by Mr. S. A. Frye, business manager at Otterbein College. The coals, of ten different suppliers, from four different states and eight different seams, represented competitive products economically available to the heating plant. The tests were conducted by the usual plant personnel which consisted of the plant superintendent and two assistants. No outside supervision of any kind was given during the tests. The tests were made from Dec. 13, 1960, through Mar. 4, 1961. During this period, 300.03 tons of coal were burned in the No. 3 unit. This tonnage represented 25.5 pct of the total coal requirements for the 1960-61 season. Operation during the period was typical for the comfort heat load which the plant supplies. Of the 78.12 operating days, the boiler was on bank or holdfire 14.31 days, or 18.32 pct of the time. The length of individual bank periods varied from 2 hr to three days. Run periods varied from about 2 hr to 7.26 days. DESCRIPTION OF PLANT AND HEATING EQUIPMENT The new Otterbein College heating plant, completed in the summer of 1957, is located in a 50x50 ft boiler room on the lower level of a combination maintenance building and heating plant. The building is situated away from the main college complex on a hillside, which allows for on-grade delivery of coal to the three bunkers. The bunkers are located directly above and connected with the individual stoker hoppers. Fig. 1 shows an in-plant view of the Coal-Pak installations. The No. 3, hot water, unit is at the extreme right. The heating plant consists of three 200-hp Coal-Paks. Unit No. 1 is operated at 4 psi and carries the steam heating load of 14,000 sq ft EDR (steam) installed in two buildings. Provision has been made to convert these buildings to hot water heating at some future date. Since the unit under maximum load conditions operates only at about 50 pct steaming capacity, a heat exchanger was installed to convert steam to hot water. Thus, during light loads in mild weather the unit can be employed more efficiently to supply both steam and hot water and eliminate the need for operating either of the two hot water units. Units No. 2 and No. 3 serve a hot water heating circuit for a large group of 21 buildings of the College. These boilers deliver hot water at temperatures up to 200°F to the various buildings through distribution lines in tunnels. The present calculated maximum heat loss for the College is 15.195 million Btu per hr. The Coal-Paks have individual design capacities of 6.700 million Btu per hr. Table I lists the design date for the Coal-Paks. The No. 3 unit was chosen as the test unit, primarily because of its bunker location at the far end of the delivery ramp. Because of this location any excess coal above bunker capacity could be conveniently stored and isolated from the adjoining bunkers. Each bunker has a capacity of 30 to 35 tons of coal. The No. 2 and No. 3 hot water generators are equipped with automatic controls which maintain the

Jan 1, 1962

By Nestor Torres-Acuña, Fathi Habashi

Metallic copper of purity equal to commercial electrolytic copper is deposited during the anodic dissolution of technically available white metal, Cu2S, in m acidic solution of' copper(II) sulfate as electrolyte and copper sheets as cathodes. The process may be applied on commercial scale since it has the advantage of by passing the converting and poling steps, decreasing pollution problems due to SO, and recovering the sulfur of the white metal in the elemental form. The anodic dissolution is believed to take place in three steps: The working conditions that lead to 98 pct cathodic current efficiency are: 30 gpl Cu2+ and 100 gpl H2SO4 in electrolyte, current density 1.0 amp per sq dm, and tetnperature of bath 40°C'. DURING the smelting of copper, nickel, lead, and zinc sulfide ores, large amounts of sulfur dioxide are produced. Because the release of this gas in the atmosphere causes pollution problems, many attempts were made in the past for its recovery either in liquified form or processed to sulfuric acid. Although these two forms of sulfur are the usual forms that are consumed by the chemical and metallurgical industries, their storage and transport usually raise economic problems. It was realized, long ago, that a process by which sulfur can be recovered directly in the elemental form would be most attractive. Elemental sulfur is easily stored and transported, and is readily converted to SO2, H2S, or H2SO4 when needed. Processes aimed to recover the metals from their sulfide ores, and at the same time by-product sulfur in the elemental form, were recently reviewed.' As early as 1882, Marchese2 patented a process in which a matte having the composition: Cu 15, Pb 14, Fe 41, and S 25 pct was electrolyzed. According to cohen3 the process was applied on a pilot scale by the Aktiengesellschaft fur Bergbau-, Blei-, und Zinkhiitten-betriebe at Stolberg, Rheinland, Germany. The process was unsuccessful, however, owing to the contamination of the electrolyte during electrolysis. Studies in this field were reported by Bernfeld4 and Egli.5 Borchers et a1.6-8 developed the process further by using white metal (Cu 78.2 pct, S 19.6 pct) instead of the matte, and a pilot plant was operated by the Mansfeldschen Kupferschiefer Co. at Eisleben, German~.~ It appears, however, that the process never went into commercial operation. Further studies in this direction were carried out by Russian workers.10-23 The International Nickel Co. of Canada erected a pilot plant at Port Colborne in 1951 to apply the process to nickel matte which is essentially pure nickel sulfide."'- 26 Full-scale plant went into operation at the Thompson Refinery in Manitoba in 1964.27 In this method nickel sulfide matte is melted and cast into anodes analyzing 76 pct Ni and 20 pct S with small amounts of copper, cobalt, and iron. These anodes are then electrolyzed in an aqueous solution of NiS04 at pH 4. Further work on this method was reported in the literature.28-31 In this paper, experiments on the anodic dissolution of white metal in aqueous copper sulfate solution are reported. EXPERIMENTAL White metal, Table I, obtained from The Anaconda Co., was cast in form of anodes 10 by 7.5 by 1 cm, Fig. 1. The anodes are mechanically strong and do not break easily; when polished they gave a shiny gray metallic appearance. From the chemical composition shown in Table I, it can be seen that the atomic ratio Cu:S corresponds approximately to the formula CU2.06S. The copper in excess to the formula Cu2S is most probably present as a separate phase since metallic copper could be seen under the microscope, Fig. 2. Fig. 3 shows the X-ray diffraction pattern of white metal which matches fairly well with that of reagent-grade Cu2S.

Jan 1, 1969

By H. Foster Bain

Steel for any large structure must be imported, the Hanyang works being entirely unable to supply local demand. The United States Steel Products Co. has warehouses and small stocks at Shanghai and at normal times there are heavy importations from Europe. It is difficult to obtain any accurate index of prices, but comparison of customs valuations with Pittsburgh quotations shows angles, bars and plates to have averaged $10 gold per ton higher in China in 1914, while on galvanized and tinned plates the excess was- $38 per ton. The margin since has been greater. Previous to the war little machinery was manufactured in China. There were shipyards at Shanghai and at Hong Kong which built a small number of engines, and small foundry and repair works are found at a number of ports. Their capacity was limited and the difficulties and delays incident to transport in China so numerous that in practice every important mine has a small foundry, boiler works, and machine shop. This must be taken into account in plans for work in the country. While skilled workmen can be found it is to be noted that most of them have been trained

Jan 1, 1921

By H. M. Otte, A. G. Crocker

SLIP, twinning, stacking faults, and precipitates on well-defined planes in a crystal produce traces that are visible on either a polished or an etched surface. The purpose of this note is to establish the usefulness and limitations of the information obtained from such traces when determining the orientation of single crystals or grains in a polycrystal. Attention will be confined specifically to traces produced by octahedral planes in cubic crystals, though most of the remarks can be taken as more generally valid. The problem of determining the crystal orientation is completely solved in principle' and in particular for the special case under consideration.1'2 Tables are available3 in which the orientations can simply be looked up when given the two or three angles between the traces. Thus a quick and simple method is available for obtaining the orientation of a crystal from surface traces. In this note two examples will be given from recently published papers and will show the use of the method with the aid of the Tables, as well as some of the precautions that should be taken in the procedure. As part of an analysis of X-ray Kossel lines, Peters and Ogilvie 4 determined from them the crystallographic orientation of the specimen. They chose a Cu-5.10 wt pct Si alloy that had been partly transformed from the a (fee) phase to the k (hcp) phase. The latter formed as very straight platelets along the octahedral planes of the a phase, and a photograph (Fig. 7 of Ref. 4) was reproduced in the paper4 of the region for which the orientation had been determined from the Kossel lines. Apart from a very small area (about 0.15 mm diam) near the top center of the photograph, the region shown was clearly all of one orientation. Traces in four different directions were readily identified. Starting with the smallest angle and proceeding counterclockwise, the angles between the traces were measured to be 27° 43', 55° 16', 64 3 43 1, and32"26'. These add up to 1802 08 1, indicating a cumulative error of 8'. The correct values should, of course, be taken from the original glass plates, or at least from prints on distortion-free paper, so the precision of the quoted values is actually greater than their accuracy. The solutions can be looked up in the Tables3 in four ways, depending on how the angles are paired, tion for the minutes of arc is required and the solution obtained will be identical for all the four pairs if the values of the angles are not in error. In a practical situation, such as this one, there are several sources of error, so that the four solutions obtained will all be slightly different. An arithmetic average of the four solutions is mathematically incorrect, but may, in special cases, give an answer close to the correct one. A rigorous procedure for refining the orientation has been described by Mackenzie;5 he gives a least-squares method that can be applied, if necessary, to this type of problem. The orientation thus determined can be compared with the one from the Kosel lines and discrepancies larger than the experimental error assigned to nonor thogonality of the X-ray or electron beam to the specimen. Here the example will only be used to illustrate the general procedure for making a very rapid rough check. From measurements on the stereographic projection, Fig. 9 of the paper,4 we can determine that the surface normal is approximately (0.434, 0.704, 0.562). From the Tables3 we find, using 28°, 55" as a good approximation to the first pair (27°43', 55°16'), that the surface normal is (0.453, 0.678, 0.579).* These two values are within 2 deg of each

Jan 1, 1967

By Larry Kaufman, Joan B. Berkowitz-Mattuck, Edward V. Claugherty

The oxidation characteristics of hafnium and zirconiunr diboride were measured between 1200 and 2200'K by a thermal- conductivity method which continuously ttzeasures the rate of reaction of oxygen with the diboride and by a metallographic air oridation method zuhich provides a measure of the total arr7ount of bovide conuerted to oxide for a given time interval. The oxidized specimens obtained from the tl~eritaal-coi~ductitrity method were also examined by quantitatire metal-lographic procedares. The significant results obtained in this investigation reveal that metal-rich compositions of lzafi~iutil diboride proride the most oxidation-resistatzt material up to 2000°K; hafnium diboride is tmre oxidation- resistant titan zirconium diboride at all tempevatures examined; the morphology of the oxide formed on H/B2 and Llie temperature coefficient qi. the oxidation rate constants change at the temperature of ttze monoclinic to tetragonal phase transition] in HfO2; the oxidation of neither HfB2 nor ZrB2, results iN catastrophic Jazlure at lorc. oxygen pressures; and pvefevetztial gvaLti boundary oxidation was not obsevued for either HfBi or ZYB, A comprehensive study of the high-temperature characteristics of refractory transition-metal di-borides is currently in progress. This program has included investigation of the physical, thermal, and thermodynamic properties of TiB2, ZrB2, HfB2, NbB2, and TaB2. In addition, aspects of the synthesis and fabrication of such materials have been studied. In view of the diverse nature of this research, a number of other laboratories have actively participated and contributed specific capabilities for analysis and characterization of these materials. As a consequence, an extensive description of the relevant properties of these compounds has emerged which is central in evaluating their high-temperature (1200" to 2500°K) performance. To date, information on thermodynamic stability, specific heat, and vaporization characteristics,1 hot hardness and electrical resistivity,1, 3 therma1 expansion:'4 and thermal conductivity 1, 5 has been presented. This information has been generated on materials of the highest purity (98.5 to 99.9 wt pct Me + B) and density currently available. Samples fabricated by zone melting6 and high-pressure hot pressing"3'7 techniques have been used to generate suitable specimens for all of the aforementioned studies. dation characteristics of the most oxidation-resistant of these materials, hafnium and zirconium diboride, is presented and a description of the synthesis and the experimental procedures used to prepare and characterize specimens is given. The high-temperature range under consideration (1200" to 2200°K) and the known dependence of oxidation characteristics on sample chemistry, density, and oxidation conditions required a close coupling of the synthesis, fabrication, and evaluation procedures.8 This was accomplished by continual surveillance of chemical composition of starting materials before and after specimen fabrication and by evaluation of density, phase constitution, and microstructural features prior to and after oxidation exposure. I) PROCUREMENT AND CHARACTERIZATION OF STARTING MATERIALS In view of the current state of the art in fabricating refractory boride materials, the methods used in preparing samples for the present study are given in detail as follows: starting materials were purchased in high-purity powder form and fabricated by high-pressure hot pressing into 0.40 by 1.00 in. bars from which oxidation specimens were obtained. The hafnium diboride used in this study was purchased from Wah Chang Corp.; the zirconium diboride from U.S. Borax and Chemical Co. These powders were routinely characterized by quantitative chemical analyses for metal, boron, carbon, oxygen, nitrogen, and iron, by qualitative emission-spectrographic analysis for trace impurities, by X-ray procedures for extraneous phase identification, and by powder densitometry for comparison with X-ray (theoretical) density. Hafnium and zirconium metal and elemental boron were also purchased as high-purity powders and characterized for impurities by emission-spectrographic analyses. The hafnium diboride was procured in three shipments which were designated as HfBl.g7(1), HfB1.88(2A), and HfB2.12(2). The indicated stoichiometry is based on the atomic ratio of total boron to total hafnium; the number in parentheses identifies the shipment number. Shipment 1 was 5 1b, shipment 2A, 1 1b, and shipment 2, 8 1b. The zirconium diboride was procured as a 20-1b shipment and designated as ZrB1.89(1). A small quantity of purified zirconium diboride was also supplied and designated ZrB1, 9(P). The averaged results for chemical analyses which were generally performed according to the procedures set forth in the compilation by KrieGe9 are presented in Table I. Qualitative spectrographic analyses indicated that Ca, Cr, Ti, Si, Zr (in H~B~), and A1 were present at levels between 0.01 and 0.10 wt pct. Other metallic elements were found to be less than 0.01 wt pct. Since it is virtually impossible to purchase these materials in the desired quantities (5 to 20 lb) as single-phase compounds it is necessary to obtain

Jan 1, 1968

By D. Turnbull, J. C. Fisher

ACCORDING to the "reaction-path" modell,2 of martensite nucleation, the shear angle of the embryonic martensite plate must be treated as a variable, and included in any calculation of nucleus critical size. Also, as can be deduced from this model, the interfacial free energy between austenite and martensite does not reach its final value until the shear is completed. It is zero for zero shear angle. However, in order to account for the kinetics of the martensite transformation, some sort of interfacial energy barrier appears to be necessary even with the reaction-path model, for otherwise the volume and the energy of formation of the critical size nucleus both collapse to zero.3 Cohen independently suggested that surface energy could be incorporated into the reaction-path model, with the overall free energy of a martensite embryo being a function of its volume and shear angle.' It is possible to estimate the energy associated with the formation of a critical-size martensite nucleus starting with the reaction-path model and including a surface free-energy barrier. As the dependence of interfacial free energy upon shear angle is unknown, a simple type of dependence will be assumed, with the belief that the true dependence would not lead to appreciably different results. Consider the work required to form a lenticular martensite plate with radius r, thickness t, and shear angle 8. There are three contributions; one being the interfacial free energy, one being the free energy change in the martensite plate, and one being the free energy increase in the surrounding austenite. The interfacial free energy u is assumed to depend upon the shear angle 0 according to the relationship s=s0(?/?0)n [1] where 8, is the equilibrium shear angle and n is an exponent that may lie in the range 0 n 2. The work required to form the interfaces of a martensite plate then is W. = 2pr² s0(?/?0)n [2] The free energy change per unit volume of martensite is composed of two parts, one the ordinary volume free energy ?f1. which is negative, and the other the elastic strain energy G?m²/2, where G is the shear modulus and 7, the shear strain relative to the martensite structure. This expression for the strain energy is valid only when the shear strain ym, is sufficiently small that the martensite is within its linear elastic range. There is no doubt that ym, lies beyond the linear elastic range for embryos that are considerably subcritical. However, for critical nuclei it will be shown that ym, is 1.5 pct or less, within the linear elastic range of martensite. For embryos of nearly critical size, then, the strain energy of the martensite is correctly given by G?m²/2. The shear strain in the martensite is ym, = 8, — 8, and the work required to form the strained martensite is Wm --= (pr²t/2) [?fv + G(?O - ?)²/2] [3] The free energy change in the austenite is entirely that due to elastic distortion. The elastic strain is not uniformly distributed in the austenite, being large near the martensite plate and small elsewhere. Approximately, however, the energy corresponds to a uniform shear strain ya= (?t/2)/r [4] throughout the volume 4pr³/3 surrounding the plate. The work required to strain the surrounding austenite then is Wa = (4pr³/3) (G?a²/2) = (G?²/6) prt² [51 For simplicity, the same shear modulus G is assumed for each structure. The total free energy for forming a plate then is W = W3 + Wm + Wa. = 2pr² s0 (?/p?0)n + (pr²t/2) [?fr+G(?0-?)²/2] + (G?²6) prt2 [6] This expression is correct for nuclei and for embryos of nearly critical size, where, as will be shown, the strain energy in the martensite is correctly given by the expression G (? — ?)². Having W as a function of r, t, and 8, as in Eq. 6, there is a saddle-point where W has a stationary value, W subsequently decreasing indefinitely as the nucleus volume increases along the reaction path. The stationary value of W is the energy of the critical nucleus. The critical nucleus has radius, thickness, and shear angle such that ?W/?r - awlat: = ?W/?p? = 0. Performing these differentiations and calculating the critical nucleus energy, W* = [8192p(G?/6)²;s/27 ?fv4] [7] where a= (?/?0)3n+1[l +G(8"-8)'/2af.]' [7a] and where 8 is to be determined from the equation (1 + 3n/4) + G8(6O - (9)/[Af. +G(6>o-6>)72] = 0 [8] For ?f, near —200 cal per mol or —10" ergs per cc, and 8, near 1/6, as for iron-base alloys, Eq. 8 gives ?0 - ? ~ - (4 + 3n) ?f1./4G0O [9] as the difference between the equilibrium shear angle and the actual shear angle for a critical nu-

Jan 1, 1954

By R. W. Smith, R. W. Lai

Very recently there has been a quickening of research effort in the field of silicate mineral flotation, particularly as regards the flotation of beryl.' These studies have already contributed much to the understanding of both anionic and cat-ionic flotation of silicates. Of particular interest concerning anionic flotation has been the work done on heavy metal activation of silicates by M. C. Fuerstenau and co-workers.4,7 In addition. in the paper by Fuerstenau, Rice, Somasundaran and Fuerstenau7 attention was paid to low pH anionic flotation of beryl in the absence of activators. This subject was also covered in the papers by Avotins,3

Jan 1, 1967

By F. Gordon Smith

ALTHOUGH several geological indicators of the critical type are known, including quartz inversions and decomposition of hydrous minerals such as serpentine, there are very few of the general type. Solid solutions are excepted, but the limitations of use are very restricted and interpretations are sometimes ambiguous. General methods for determining temperature and pressure conditions during the crystallization of minerals would have considerable scientific and economic value. It is not the purpose of this paper .to, discuss the various methods of geological thermometry and barometry, but to present one general method, applicable to all minerals, and to describe what progress has been made in the methods of measurement. The general method, in. brief, is a study of the stress conditions in and around various types of foreign inclusions which are trapped in minerals during growth. The method depends upon the fact that any homogeneous gas, liquid, or solid will in general have coefficients of thermal expansion and volume compressibility different from those of any given mineral. Therefore stress must develop in and around all types of inclusions in minerals if the temperature or the pressure, or both, are changed from the conditions which prevailed during the growth of the mineral. The methods of measurement consist of determinations of the temperature-pressure conditions of fit of the inclusions in the host mineral. The types of inclusions in minerals are: 1-gas, or liquid plus vapor, when observed at room temperature, due to crystallization under pneumatolytic conditions; 2-liquid, or liquid plus vapor when observed at room temperature, due to crystallization under hydrothermal conditions; 3-glassy solid, or devitrified glass, due to crystallization under magmatic or high temperature metamorphic conditions, in a siliceous liquid; and 4-crystals, due to overgrowth of other minerals crystallizing simultaneously or of other minerals which crystallized previously. A survey of the literature shows that much valuable earlier work on inclusions, especially that carried out in England in the last century, has dropped out of current knowledge. The following is a brief summary of the significant, contributions to the problem up to the present day. Davy in 1822 asserted that fluid inclusions in minerals consist of an aqueous solution of salts and a gas bubble, the whole being either at lower or higher pressure than atmospheric.1 At intervals from 1823 to 1862, Brewster contributed information concerning other types of inclusion consisting of 1-aqueous solution, a much more expansible liquid, and a gas; 2-aqueous solution, salt crystals, and a gas; and 3-the very expansible liquid and a gas. The very expansible liquid fills the gas space between 20° and 30°C. Compression strain effects were seen around inclusions in diamond, topaz, and other minerals.2-6 Sorby in 1858 and 1869 further advanced the study begun by Davy, stating that fluid inclusions represent a sample of the mother liquor of crystallization and that the degree of filling of aqueous inclusions at room temperature defines the temperature-pressure relations during formation of the host. The degree of filling may be measured by determining the minimum temperature of filling of the inclusion by the liquid phase. The very expansible liquid in some fluid inclusions is liquid. carbon dioxide. The temperature at which salt crystals in fluid inclusions completely dissolve in the fluid is the minimum temperature of formation. Inclusions of glass or devitrified glass indicate crystallization from a melt. Inclusions of crystals in minerals are often centers of strain, which may be seen by optical effects or by radial tension cracks. Sorby realized that an analysis of stress-strain relations about inclusions could be used to provide precise data on the temperature of crystallization, but the matter was never pursued.7,8 Hartley (1876, 1877),9,10 Hawes (1881),11 Wright (1881),12 Johnsen (1920),13 and Holden (1925)14

Jan 1, 1952

From "Knots and Lashings," of Oct. 10, 1918, the military publication of Canadian Engineers' Training Depot, St. Johns, Que. During the early hours of Sunday morning last (the 6th inst.) there passed from our midst one of the most affable and estimable officers who, during our brief connection with the Depob, it has been our fortune to meet. Sidney A. Lang was born in Ontario in the year 1884 and lived in Toronto for a number of years. He ahtended Toronto University, where he displayed remarkable ability, and after graduating followed the profession of mining engineer. Being young and endowed with the spirit of adventure, he followed his profession in the mines of Chile and Peru; he could speak Spanish fluently. Always a reserved man, of refined tastes, it was difficult to induce him to relate some of his adventures in those countries so shrouded in mystery and romance. He joined the Canadian Engineers in Toronto as a sapper and shortly after came to this Depot.

Jan 1, 1920

By F. F. McINTOSH

STEEL is a general name applied to the alloys of iron and carbon. These alloys always contain , other elements such as manganese, silicon, sulfur, and phosphorus. Manganese and silicon are usually considered to be alloying elements. Sulfur and phosphorus are regarded as impurities. Tradition has damned them both and, in addition to classifying them as impurities, has established the general impression that many of the ills of steels are the result of the presence of one or both of these elements. Metallurgists are inclined to question the soundness of this tradition. The fact that one blast-furnace plant is devoted entirely to the production of ferro-phosphorus which is used to rephosphorize basic open-hearth steel, and that sulfur is added to many heats of steel, indicates that these so- called impurities might also be classed as alloying elements, and that the tradition which damns them may not be altogether fair and well founded. A critical search through the technical literature on the effect of phosphorus on steel will disclose many references to its harmful effect. Most of these are general in character, but there has been a strong tendency to accept this traditional belief in the harmful effect of the element.

Jan 1, 1926

The curve of Fig. 3 for pyrite requires modification. There 11% a range of pH values from 6.2 to 7.8 for which no cyanide is required to prevent contact. From 7.8 to 11.3 a small amount of cyanide is required, with a sharp maximum of 20 mg. per liter at pH=9.6. The mistake arose from considering too few solution compositions (17 were considered) in constructing the curve. Subsequent work with other collectors indicates that the corrected curve is typical for such systems; it illustrates the further development of the dip in the pyrite curve of Fig. 5.

Jan 1, 1938

By Fred C., Bond

DETERMINATIONS of circulating load in a classifier operating in closed circuit with a grinding mill are usually made from screen analyses of the mill feed and classifier products. The percentage remaining on someone selected screen size must be used in the calculation, and, because of the wide variation in average particle size between the mill feed and classifier overflow, an accurate and representative result at the selected mesh is often difficult to obtain. To overcome this difficulty and reduce the time required in making the determination, I have devised a method which uses the liquid-solid relationship in the circuit rather than screen analyses. The quantities required for the calculation include only the usual dilution and feed tonnage measurements, which should be made regularly in the course of ordinary mill operation and, in addition, the amount of new water added to the classifier, or to the head of the mill with the new feed. If measurements are obtained on both of these quantities it is possible to check the calculated circulating load, and to calculate the percentage of moisture in the new feed to the mill; or, if the latter is known, to check the feed tonnage.

Jan 1, 1937

By C. W. Dunham

FOR discharging and diffusing the gases from the reverberatory furnaces and converters the Morenci Reduction Works has been provided with one of the largest reinforced concrete chimneys ever built. It is 612 ft. high from the bottom of the foundation to the cap at the top, and has a minimum diameter of 24 ft. inside the lining near the top and 38 ft. 10 in. at the bottom. Its estimated capacity, at the elevation of 4338 ft. above sea level, is about 1,000,000 cu.ft. of gas per minute at a temperature of 500° F. and an atmospheric temperature of 90°. Under these conditions the draft will be at least 3 in. of water at the flue inlet. The gases enter the chimney through a steel balloon flue 33 ft. high and 15 ft. wide, the bottom of which is 19 ft. above the ground.

Jan 1, 1942

By AIME AIME

MILLING called for four sessions and a luncheon and covered broad ranges from speculative theory to basic practice, and from coal to gold. An attractive and profitable feature was the "get-together" of "coal preparers" and "ore dressers" for mutual discussion of their common problems. The first event was the Milling Luncheon in the Engi¬neers' Club on Monday. A concentrated group of high¬grade appeared and many matters were discussed, but only two called for official action: The meeting unanimously recommended the use of the Tyler standard sieve scale as against any new scale, and also the adoption of micron designation for fine particles in place of decimals of an inch or millimeter, or theoretical mesh size. The real sessions began on Monday afternoon with a menu of Mineral Physics for the ore dresser, metallurgist, geologist, physicist, et al. Four papers by the staff of the U. S. Bureau of Mines dealt with the chemical and physical aspects of mineral structure as revealed by the first two years of research to date. A fifth paper by Dr. M. J. Buerger correlated to the foregoing the knowledge of "The Lineage Structure of Crystals" and finally a sixth paper was a progress report of further developments of the last year in explosive shattering of minerals by the U. S. Bureau of Mines.

Jan 1, 1933

By Eckley B. Coxe

AT the first meeting of the Institute, a paper was read by Mr. Rothwell, calling attention to the importance of at once considering the great waste of anthracite coal under the present system of mining and preparing that fuel for market, and a committee was appointed to examine and report upon the method now in use. About the time the committee were selected, a general resumption of work, after the long strike, took place in the coal regions, and the members of the committee have not in consequence had as much leisure to devote to the study of the question as its importance de¬mands. In the present preliminary report we shall merely sketch out the plan which we propose to pursue in our investigations, so that when our fellow-members are called upon to assist us in our work by fur¬nishing statistics, etc., they may understand more clearly for what object our inquiries will be made. As one of the most practicable methods of obtaining the informa¬tion requisite for an intelligent consideration of our subject, we pro¬pose to send, from time to time, to members and others interested in the mining of anthracite coal, printed sets of questions relating to points under consideration. We would here impress upon our fellow¬-members the importance of filling up these papers as far as they can from their own knowledge. We would, of course, wish each ques¬tion to be answered as fully as possible, but we do not expect every one to answer them all. One question well answered would be of more use to us than the whole paper filled up with answers, many of

Jan 1, 1873

By D. J. Brown

Froth flotation is now used to recover millions of tons of fine coal,(-1/2 mm), every year. For example, in Great Britain, 4-million tons of flotation concentrate was produced in 1960, about 3% of the mechanically cleaned out- put and 2% of the total salable output. Coal flotation is not so widely practiced in the U. S. as it is in Europe since much American coal occurs in thick seams of high grade and does not require cleaning. However, in the U. S., as elsewhere, the increasing mechanization of coal mining machinery has lead to a lowering in size and grade of product from the mine and to an increasing need for cleaning of coal to improve the grade before sale. With an increasing proportion of fines in the product, froth flotation has assumed a greater importance as a cleaning process for coal. In the last three years, over 40 flotation plants have been put into operation in the Appalachian area of the U. S.80 Fine coal recovered by means of froth flotation may be used as a powdered fuel alone or mixed with coarser grades. The removal of this fine coal from washery water makes possible the recirculation and re-use of the water and also leads to improved and more consistent operation of jigs and other separating machines which function less effectively when there is a build-up of fine solids in the washery water. Finally, froth flotation is of assistance in the production of a clarified effluent which may be discharged from the washery without polluting local rivers, streams, or wells. The 1961 cost of installing a new coal flotation plant in Great Britain would be about $19,000 per ton of dry solids input per hour, including the cost of new buildings and coal and shale filtration and water clarification equipment. The figure for the gravity concentration of the large coal, the main plant, would be about $4,800 per tph. The operating cost for the new coal flotation plant would be about $1.10 per ton. Even at this level of cost the cleaning operation is economic in Great Britain for coking coals. In the U. S., a froth flotation cost of 14$per ton of cleanfilter cake was quoted early in 1958 for one operation,l though admittedly this was a special case. Assuming a cake moisture con- tent of 30%, it is estimated that this cost is equivalent to a figure of 256 per ton of dry solids input, for the incorporation of coal flotation in an existing washery. A cost oh 416 per ton of concentrate was quoted76 in a recent analysis of flotation economics, this cost being brought up to $1.11 per ton after payment of royalty, welfare and association dues and taxes. Capital cost was $120,000 for a plant producing 133 tpd of concentrate. Before considering the treatment of coal by flotation it is appropriate to survey some of the properties of coal which are of importance in the process.

Jan 1, 1962

By John V. Beall

During the last quarter of 1948, two new machines, which may revolutionize the coal mining industry, made their first public appearance within two months of each other. Both are designed to mine and load coal continuously. On Oct. 27, the "Colmol," invented by C. H. Snyder and A. E. Lamm, president and executive vice-president, respectively, of the Sunnyhill Coal Co., was demonstrated at the No. 8 mine of that Company at New Lexington, Ohio. The "Joy Continuous Miner," invented by H. F. Silver, president of the Silver Engineering Works, Denver, and developed and manufactured by the Joy Manufacturing Co., was demonstrated at the Mathies mine, near Finleyville, Pa., in the Pittsburgh coal seam.

Jan 1, 1949

By David R. Maneval, H. B. Charmbury

At the turn of the century, iron and coal were the keys to industrial prosperity. At that time, Pennsylvania was the leading mineral producer in the Country, producing 200,000,000 tons of coal in a typical post-World War I year. Close to 90% of the Nation's energy was generated by coal. Today, coal accounts for but 22% of the energy generated, leading to a drastic reduction in Pennsylvania coal output. In contrast with the post-World War I days, 1964 figures showed a drop in production to 93,846,480 net tons. This huge decrease has created a serious economic problem in a state so much dependent on the coal industry. Along with the reduction in coal production, employment has dropped considerably through these years (the number of miners alone has gone from 280,011 to 43,149), thus causing a heavy burden of supporting the unemployed on the taxpayers. As far back as 1939, coal operators and economists saw the problems on the horizon and the need for improvement. Realizing that research was a way to improve and stabilize the industry they demanded legislative action. The Legislature reciprocated with a small grant to The Pennsylvania State University with the provision that each State research dollar be backed with one from the industry. This initial grant laid the foundation for the further growth in coal research.

Jan 3, 1966

By H. DeWitt Smith, W. H. Sirdevan

The mine operated by the United Verde Copper Co. is situated near Jerome, Ariz., on the eastern flank of the Black Hills, at an elevation of approximately 5500 ft. (1676 m.) above sea level. The mine and the town of Jerome are located on a steep hillside, which slopes rapidly to the valley of the Verde River. The smelter town of Clarkdale is situated in the Verde valley 4.1 mi. (6.5 km.) distant in an air line from Jerome, and at an elevation of 3560 ft. The two towns are connected by the standard-gage railroad of the Verde Tunnel & Smelter R. R., a subsidiary of the United Verde Copper Co. This road is 11 mi. long and has a compensated 4 per cent. grade. At Clarkdale, it connects with the Verde Valley branch of the Santa Fe R. R. The narrow-gage (36-in.) line of the United Verde & Pacific Ry. was, until 1919, Jerome's only rail connection with a main trunk line. This road connects with a branch line of the Santa Fe system at Jerome Junction, after traversing 26 mi. of mountainous country. From the time of its location, in 1876, until its purchase by Senator W. A. Clark, of Montana, in 1889, the United Verde mine was worked on a small scale and shipments of high-grade gold-silver ore were made from the surface workings. Until 1894, when the narrow-gage railroad to Jerome Junction was completed, all supplies for the mine and smelter

Jan 1, 1922

By Jean A. Andre

In 1969 "Société des Mines et Fonderies de Zinc de la Vieille-Montagne" produced 221,000 tons zinc ingots and 22,000 tons zinc dust, thus rating highest on the world's zinc producer list. The Company's plant at Balen, Belgium, with an output of 135,000 tons electrolytic zinc is the most important one of this group; it is an integrated metallurgical zinc and lead unit with recovery of most by-products. The present paper describes the zinc production facilities. The roasting section includes four fluid bed roasters (total concentrates charged: 740 tpd) and two sulfuric acid having a corresponding capacity. Calcine is leached in a fully automated plant by a two-step continuous process. Leaching residues, mixed with lead furnace slag, are treated in a shaft furnace. The recovered zinc fume is leached separately. In the near future, the leaching residues are to be treated by a wet process, developed in the Company's own laboratories. Purification of solutions is accomplished in two steps: antimony and zinc dust are added in the first one, only zinc dust in the second one for removal of cadmium. Electrolysis cells are partly of a conventional type. A new type of cell configuration was put onstream in October, 1969, planned output 60,000 tpy. Main features of the new cell rooms are the fully automatic pulling, stripping and replacing of the jumbo-size cathodes (immersed area 2.6 m2). Zinc cathodes are remelted in induction furnaces followed by in-line automatic casting and stacking machines.

Jan 1, 1970