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By W. A. Hockings, M. E. Volin, A. L. Mular

Batch grinding tests at short times were made in a laboratory rod mill with 10 x 14 mesh quartzite in corn syrup-water mixtures of varying viscosity. The weight fraction broken and size modulus were found to be independent of viscosity up to 20 cp, but at higher viscosities breakage of the feed decreased while the size modulus increased. The mechanisms of the effect of viscosity are discussed on the basis of particle dynamics. The effect of viscosity of the suspending medium on grinding has not been accorded much attention, although both wet and dry grinding are used in bene-ficiation processes and the viscosities of water and air differ greatly. schweyerl investigated the rates of grinding of quartzite in a pebble mill with air, water and glycerol as the suspending media and found the rate in terms of new surface developed per thousand revolutions of the mill to be constant and dependent on viscosity up to 20,000 rev, but to decrease and become independent of viscosity for longer grinding times. Based on considerations of the drag forces exerted on suspended particles and on the grinding media under turbulent and laminar flow conditions, viscosity can be expected to inhibit the rate of grinding in a given system as it changes the flow conditions from turbulent to laminar. The objective of this study was to determine the effect of viscosity of the suspending fluid in the batch grinding of a homogeneous feed of uniform size for short residence times, a region of more practical interest than the long times investigated by Schweyer. In actual grinding operations, the viscosity of the fluid does not change, but the consistency (apparent viscosity) of the pulp increases with additions of fines. In this study the consistency of the pulp was not measured because it was thought that the grinding times were too short to alter the apparent viscosity appreciably.* *Minus 200 mesh quartz at 65% solids has an apparent viscosity of about 25 cp as measured by a simple consistometer. Personal communication from D. F. Kelsall to A. L. Mular. For the purposes of this study, the effect of viscosity on grinding can be shown sufficiently well on the basis of the cumulative weight fraction finer than the feed size at any time, and the size modulus, ku, of the finer than feed sizes. The parameter, ku, is obtained from a form of the Gaudin-Schuhmann equation. METHOD AND MATERIAL The grinding tests were accomplished in a laboratory rod mill 10-1/2 in. long and 8 in. bore with a 20-lb rod charge consisting of two 1-in., eight 3/4-in., eight 1/2-in., and eight 1/4-in. diam rods. The mill speed was 43.2 rpm. Charges of 400 g 10 x 14 mesh Wisconsin quartzite** were ground for times of 60, **Courtesy of Minnesota Mining and Mfg. Co. 120, 180 and 240 sec in 500 cc of fluid made up of corn syrup and water in measured proportions. The temperature of the pulp was measured immediately upon completion of grinding, and the size distribution of the quartzite was determined by a wash-wet-dry screening technique. Corn syrup was chosen because of its high viscosity and ideal viscous behavior, and also because its density is not greatly different from that of water. The viscosity of each fluid mixture was measured with an Ostwald viscometer at 25 C. and a correction was made for the temperature of the pulp according to tabulated values for sucrose solutions. The density differences were so small that they probably did not significantly affect breakage. RESULTS AND DISCUSSION The conditions of the grinding tests, computed viscosities, weight fractions finer than the feed size and values of a and ku are listed in Table I. The values of a and ku were taken from the linear portions of the log-log plots of cumulative weight fraction finer vs size for the various grinding times. The plots for tests in fluids of average viscosity 8.6 cp are shown in Fig. 1. In general all of the sets of curves were linear throughout most of the size

Jan 1, 1965

By Michael Hoch, Walter C. Wyder

The Cr-Ni system was investigated between 900" and 1400°C and 45 to 100 wt pct Cr. The Cr-Fe system was studied between 1200" and 1385°C and 70 to 100 wt pct Cr. Powdered alloy compacts were produced from high-purity chromium, nickel, and iron poulders. The compacts were sintered in an argon atmosphere and then machined into small, 0.030-in. diam cylindrical specimens. These were then heated by induction in the high-temperature X-ray diJfractwn camem under a helium atmosphere and irradiated with copper Ka radiation. From the results of the high-temperature X-ray diffraction patterns it can be concluded that chromium exists only in the bcc form, and that no eutectoid reaction occurs in the Cr-Ni system in the temperature interval 900" to 1400°C. The X-ray diffraction patterns are in agreement with the Cr-Ni phase diagram of Bechtoldt and Vacher, and Williams. It is possible that chromium undergoes one or more allotropic transformations between room temperature and its melting point. Most of the evidence was gathered from the study of the chromium-rich ends of the Cr-Ni and Cr-Fe binary systems. All the investigations rely heavily on indirect methods: thermal analysis, electrical resistivity, and micro-graphic inspection of quenched samples. In the present work the chromium-rich side of the Cr-Ni and Cr-Fe binary systems is studied, using high-temperature X-ray diffraction techniques. Thus the coexistent structures can be observed directly at temperature. PREVIOUS INVESTIGATIONS ~ansenl shows two possible Cr-Ni phase diagrams 1) a simple eutectic [Liquid =a-Cr(bcc) + y-Ni(fcc)], 2) a eutectic [Liquid+ P-Cr(fcc) + y-Ni(fcc)] and euctectoid [p-Cr(fcc) ==a-Cr(bcc) + yNi(fcc)]. The latter is based on the work of Grant and co-workers.2 Stein and rant'" placed the eutectoid composition at 1215°C and 68 wt pct Cr; their diagram is shown in Fig. l. Similarly, Sully3 also discusses the two proposed diagrams. Price and Grant,4 working on an isothermal section of the Cr-Ni- Fe ternary diagram at 1300°C, made no mention of a ß-chromium phase (fcc). The ß-chromium structure was also reported by Misencik 5 in the Cr-Nb(Cb) system. Williams 6 studied the Cr-Ni system using precipitation techniques. His work included the chromium and nickel solvus lines below 1250°C. He reported no ß structure nor any eutectoid reaction. The solvus lines which he produced were very similar to those of Jette, et al 7 below 1000°C. He concluded that Grant and coworkers2 had misinterpreted their data, and that effects which they had reported were actually due to the rapid and extensive precipitation of the nickel-rich phase. Bechtoldt and vacher8 used sintered powdered alloy compacts in an attempt to find the ß phase. They used metallographic and room-temperature X-ray diffraction techniques and also failed to detect the fcc chromium phase. They attributed the low slope of the chromium solvus at approximately 1150°C as the probable cause of the abundant precipitation obtained. They also stated that the precipitation was the probable cause for the abrupt discontinuities in the thermal analysis and electrical resistivity data which led to the conclusion that a eutectoid reaction existed. Fig. 1 also shows their diagram. Grigor'ev and coworkers9 investigated the polymorphic changes of chromium in the Cr-Ni phase diagram. Using standard microstructural and thermal analysis techniques, they published a phase diagram which contained five polymorphic forms of pure chromium and four eutectoid reactions. They stated that the room-temperature bcc chromium transformed to a fcc structure at 930°C, to bcc at 1300°C, to hcp at 1650oC, and then finally to bcc at 1830°C. The eutectoid reactions were placed at 850°, 960o, 1140°, and 1220°C. Also to be noted was the fact that the uppermost phase, the bcc above 1830°C, was different from the upper phase of Grant and coworkers,2 the fcc above 1840°C. Their eutectoid reactions, temperatures and compositions fall

Jan 1, 1963

By F. W. Archibald

The American Smelting and Refining Co. has standardized its copper converting practice to attain a maximum unit blister production with a minimum of refractory consumption by careful location of the tuyeres and by applying magnetite coatings on the hard-burned magnesite brick linings. THE American Smelting and Refining Co. operates four primary copper smelters in the United States with a total of 17 Peirce-Smith type converters; 15 of them are 13 ft in diameter by 30 ft long, and two are smaller. Some details of operations vary with locale; however, fundamentals of design, operation, and maintenance are common to all plants. All converter shells are l-in. thick except for one new converter with riding rings on the ends which has a shell thickness of 1 1/2 in. More tuyeres can be installed with rings on the ends and hence more air can be used. Welded construction is replacing riveted. Minor shell repairs are made after each campaign. Principal causes of complete shell replacement are warpage and cracking resulting from localized over-heating. A 13x30 ft shell is being replaced after 34 years of operation and a total production of approximately 750,000 tons of blister. Converters are driven by 80 hp DC motors through company-designed worm gear reducers. Rotation is 0.38 rpm which permits the skimmer to spot the converter quite accurately for skimming slag. At three of the plants, protection against unscheduled air or power failures is afforded by the installation of emergency drives, consisting of auxiliary air motors or storage batteries. In order to avoid excessive splash out of the converters, the converter mouths are located as far back on the shells as existing flue facilities will permit. At one plant, the back of the mouth is only 13" to the rear of the vertical center line of the converter,. whereas it is 28" at another plant. Newly-lined mouth areas vary from 36 to 44 sq ft with effective operating areas about 25 pct less. It is important to keep the converter mouths as clean as possible. Dirty mouths create back pressures in the converters and as a consequence the tuyere air volumes are reduced. Mouths are generally cleaned by bumping with an empty ladle. Small mouth-cleaning rams have been used but unless extreme care is exercised the brickwork may be damaged. After considerable experimentation, the plants standardized tuyere elevations at 4 to 4% in. below horizontal center line at the shell with a downward pitch of about 13/16 in. per ft of length. Tuyeres are spaced at 6 in. centers except at the riding rings where there are no tuyeres. With this tuyere location, several of the plants now freeze magnetite slag in the bottoms, fronts, and backs up to the bottom of the tuyeres to control the internal shape of the converter. This has the effect of improving the agitation and mixing so that there is a marked increase in converting speed besides affording protection to the brick lining. Currently, the trend is to increase tuyere diameters from 1% to 2 in. to increase the air flow. At present, all converters are hand-punched but one converter is being equipped with a set of mechanical tuyere punchers. Rods for punching are % in. hexagonal smelter bar upset to ll/s in. and rods for cleaning tuyeres are upset to about 1% in., or larger, depending upon the tuyere diameter. Two of the plants use pneumatic reamers for cleaning the tuyeres between charges to minimize disturbing the coating on the inside of the refractory lining. Most of the puncher's platforms are pneumatically or hydraulically mounted so that a convenient punching position can be maintained regardless of tuyere position. For normal operations, tuyere air pressures vary from 15 to 13.5 psi, although some cycles such as magnetiting require pressures down to 8 psi. Air requirements vary from an average of 25,000 cfm on the newer installations down to 12,000 cfm on the older ones. Converter hoods are designed to protect the punchers from sparks, splash or hood accretions as well as to prevent the escape of objectionable

Jan 1, 1955

By Y. Kondo, S. Yamazaki, Z. Asaki

Thermal deco7nposition of Pyrite particles in a fluidized bed with inert gas stream was studied. Assuming that heat transfer from the surroundings to the fluidized particles controls the overall decomposition rate, rate equations for the batch process and for the continuous process were derived. In the batch experiment, a linear rate equation satisfies the experimental results and the overall heat transfer coefficient calculated from the rate constant agrees fairly well with that obtained by Leva.l1 For the continuous process, two rate equations were derived, one on the assumption of complete mixing of particles and another on the upward piston flow of particles in a fluidized bed. The former holds for a bed containing a higher fraction of decomposed pyrite realized at lower feeding rates. The latter can be applied for a bed at higher feeding rates. Thus, segregation of particles in the fluidized bed was indicated at higher feeding rates. Bed temperatures also correspond to these conditions. ThERMAL decomposition of pyrite may be represented by Eq. [I]. The pressure of diatomic sulfur gas reaches 1 atm at about 690°C. The thermodynamics,' kinetics,2'3 composition, and properties3-5 of decomposed products of such a reaction have been studied. Pyrite is a very common sul-fide mineral and is often accompanied with other sul-fides. It is of basic interest in nonferrous metallurgy to clarify the behavior of pyrite in the pyrometallur-gical processes of sulfide minerals of metals such as copper, lead, zinc, nickel, and so forth. Interest in this reaction increased recently because of possible elimination of arsenic from pyrite in processing highly purified iron oxide pellets. Producing elemental sulfur from pyrite, instead of sulfuric acid, also aroused interest in this reaction. It is indicated that the thermal decomposition of solid particles, such as calcium carbonate, proceeds through three major sequential steps: heat transfer, interfacial chemical reaction, and mass transfer.637 It is known that the decomposed product of pyrite is very porous2, 3 and the diatomic sulfur gas evolved can easily escape through this layer of decomposed product. It depends upon the circumstances, therefore, whether the heat transfer to the interface within particles or the chemical reaction at the interface determines the overall decomposition rate. The enthalpy change in the decomposition of pyrite is about 33 kcal per mole FeS2 which is comparable to that of calcium carbonate. The decomposition of calcium car- bonate becomes more and more dependent on the rate of transport of heat when reaction temperature increases, such as occurs in a fluidized bed.6'7 It is reasonable to presume, therefore, that the thermal decomposition of pyrite, an endothermic process, carried out in a fluidized bed may be analyzed according to the heat transfer controlling model. This work intends, first, to propose a mathematical model that determines the overall rate in a fluidized bed for the decomposition process and, second, to investigate a few characteristics of the fluidized bed based upon the experimental results obtained. KINETICS OF THERMAL DECOMPOSITION IN A FLUIDIZED BED It is intended in this section to obtain rate equations for thermal decomposition of pyrite in a fluidized bed by assuming that the overall rate is determined by heat transfer from the surroundings to the particles. Both batch and continuous processes are considered. 1) Batch Process. To obtain the rate equation in the batch process, the following two additional assumptions are made. First, the temperature of preheated inert gas, tg, blown into the fluidized bed is assumed to be the same as the temperature of the fluidized bed, tf. Thus, no heat exchange occurs between the gas and particles in the bed and only the heat transfer from the reactor wall kept at tw to the particles is to be considered. Second, the decomposition is assumed to start at the outer surface of the particles and to proceed toward the center. At any given time during decomposition, undecomposed pyrite remains in the tori at a temperature: td. The decomposed shell is composed of FeS1+x whose outer surface is at tp Diatomic sulfur gas evolving at the interface is heated to tf during its escape through the decomposed shell. This is illustrated in Fig. 1. With the above-mentioned assumptions of heat transfer, we have:

Jan 1, 1969

By Richard A. Queeney, Lance G. Peterson

ESTIMATES of the stacking fault free energy of copper reported in the literature show an extensive divergence of results. Based on measurements of dislocation node radii, Thornton et al.7 find the lower limit of stacking fault energy to be 60 ergs per sq cm: Jossang and coworkers4 estimate the same value should be less than 40 ergs per sq cm. Assuming stacking fault free energy to be twice the coherent twin boundary free energy, Valenzuela8 reports 72 ergs per cm, while Inman and Khan2 cite a value of 24 ergs per sq cm. which can no doubt be accounted for by almost complete absence of interstitials in solution. However, the minute concentration was sufficient to permit observable recovery of dislocation damping. Assuming a dislocation density of 1010 CM -2 and an average loop length of 0.5 , a concentration of interstitials as low as 0.01 at. ppm is sufficient to reduce dislocation damping by more than a factor of ten. The concentration of interstitial solute atoms in Ferrovac E-0.15 pct Ti was thus estimated at 0.01 at. ppm < (C + N) < 2 at. ppm. The absence at elevated temperature of strengthening due to interstitials, Fig. 1, was therefore anticipated in the titanium alloy. The temperature dependence of the flow stress below 300°K was unaffected by addition of either titanium&apos; or zirconium.&apos; Observed, however, was an initial difference in the absolute value of the flow stress below 320oK, Fig. 1. Grain size may account for the difference; the grain diameter of Ferrovac E was 30 to 45 µ and that of Ferrovac E-0.15 pct Ti 65 to 90 µ. Although second phase particles in the form of titanium carbides and nitrides precipitated in the Ferrovac E-0.15 pct Ti alloy, the spacing was too large to produce any strengthening effect. The mechanical properties measured, therefore, were inherent to the bcc lattice. 1W. C. Leslie and R. J. Sober: Trans. ASm, 1967, vol. 60, pp. 99-111. 2 H. Conrad: Journal of Metals, 1964, vol. 16, pp. 582-88. 3 A. S. Keh and W. C. Leslie: Materials Science Research. H. H. Stadelmaier and W.W.Austin, eds., vol. 1.pp. 208.50, Plenum Press, New York, 1963. 4A. S. Keh, Y. Nakada, and W. C. Leslie: Dislocation Dymmics, Rosenfield, Hahn, Bement, Jr., and Jaffe, eds., pp. 381-406, McCraw-Hill Book Co., New York, 1968. &apos;W. J. Bratina, J. T. McCrath, and H. E. Rosinger: Can. Met. Quart., in press. 6 W. J. Bratina, J. T. McGrath, and D. Mills: Suppl. Trans. Japan Inst. Metals. 1968, vol. 9, pp. 436-43. &apos;H. E. Rosinger and C. B. Craig: Can. Met. Quart., In press. %. Mills. J. T. McGrath. and W. J. Bratina: ScriptaMet,, 1968, vol. 2, pp. 311-13. The purpose of this note is to present the results of a direct measurement of the stacking fault free energy in copper. The present method avoids both the question of interaction forces of nodal dislocations and the assumption that the coherent twin boundary free energy is one-half that of a stacking fault. Fig. 1 shows a transmission electron micrograph of a stacking fault intersecting a high angle grain boundary in copper: Fig. 2 is a schematic representation of the intersecting interfaces. To produce the specimen, pure copper was rolled to 0.002 in. thickness and annealed for several hours at 875°C. These foils were strained in tension to l04 µ in. per in., and rean-nealed to 8 min. Recrystallization and/or grain growth were not observed during the second anneal, indicating that intersections such as Fig. 1 are equilibrium configurations. Many such stacking faults were observed, but few were noted as intersecting grain boundaries in well-illuminated areas. The intersection of Fig. 1 was well away from the specimen edge. Since similar defects have not been observed in fully recrystal-lized and annealed copper, it is felt that this defect is due to the tensile straining. Electron transmission

Jan 1, 1970

By H. O. Hofman

At the Lake Superior meeting, September, 1904, Messrs. A. Gibb and R. C. Philp presented a paper entitled " The Constitution of Mattes Produced in Copper-Smelting,"&apos; in which they concluded that ferrous and cuprous sulphides formed the chemical compound, 5 Cu2S, FeS. They arrived at this inference from the observed fact that a matte containing Cu,S, 90.4, and FcS, 9.6 per cent. (or, in round figures, Cu2S, 90, and FeS, 10 per cent.), did not contain any metallic copper, while one with more than 10 per cent. of FeS readily dissolved copper, and with less than this amount retained copper in mechanical suspension. Although they determined the melting-points of five mattes containing from 32.6 to 80.1 per cent. of copper, they did not draw a freezing-point curve, as the necessary apparatus was not available. A microscopical examination of 11 polished samples gave them a grayish-black field of very uniform appearance, except that five of the specimens showed the presence of metallic copper. The object of the present investigation was to supplement the above paper by drawing the freezing-point curve of the ferrous sulphide-cuprous sulphide series, and to see how the constitution could be further revealed by microscopical work. The results obtained lead to conclusions very different from those of Gibb and Philp. While their paper bears the date 1904, it was published only in 1906.2 In the meantime, Bolles3 had presented his paper, " The Concentration of Gold and Silver in Iron-Bottoms." The raw material which formed the basis of the research was a matte containing Fe, 61.68, and Cu, 11.20

Jan 1, 1908

By Gilbert E. Doan

WHEN metallurgists and other engineers meet their college classmates or former teachers, the conversation will frequently become reminiscent and finally turn to engineering education. These graduates have tested the value of their college training in the hurly-burly of the outside world so respectful attention is due their opinions, which, strangely enough, fall into two opposite categories. A member of one group expresses a deep-seated conviction that engineers should have a fuller foundation of physical chemistry, another asks for a more thorough training in mathematics, others wish a more thorough grounding in refractories, a series of courses in petroleum engineering, or in electric welding. Each alumnus of this group

Jan 1, 1937

By A. L. Moore

KAISER Bauxite Co., a subsidiary of Kaiser Aluminum & Chemical Corp., has been mining and shipping Jamaican bauxite for over a year. On Feb. 10, 1953 the first boat load of bauxite left Port Kaiser, Jamaica, for the company&apos;s alumina plant at Baton Rouge, La. To provide its own source of the aluminum ore, a Kaiser Aluminum & Chemical Corp. exploration party started initial reconnaissance in 1947. This work, and subsequent exploration, led to the acquisition of reserves sufficient to warrant a full-scale mining operation for several decades.

Jan 3, 1954

By Robert J. White

Strong&apos;s equation for calculating bottom-hole pressure reduction due to gas cutting of drilling mud is corrected, resulting in a simpler equation which is easier to use. Use of the equation is illustrated by examples. It is shown that reduction of weight of a 10 lb/gal mud by 50 per cent at the surface due to gas cutting produces only n 3 per cent decrease in hydrostatic head at the bottom of a 5,000-ft well. It is also shown that at the surface 5 volumes of air per volume of mud are required to reduce the apparent mud weight from 8.5 to 6.5 lh/gal in a 3,000-ft well. Illustrative graphs are included to show the effect of various degrees of gas cutting on apparent mud weight at depths of 2,000, .5,000 and 10,000 ft and to show variation of apparent mud weight, with depth for a 100 lb/cu ft (13.37 lb/gal) mud gas cut to 90 lb/cu ft (12.03 lb/gal) at the surface. INTRODUCTION In 1938, Strong&apos; published the following equation for calculating reduction in bottom-hole pressure due to gas cutting of drilling mud h = 1/D [p + pn/100 In p + p(1-n/100)/p(1- n-100)] where: h = depth in feet P = pressure in atmospheres at depth h due to mud column only D = hydrostatic pressure in atmospheres of a col-umn of uncut mud 1 ft high p = back pressure at well-head (in atmospheres) n/100 = fraction by volume of gas in mud at well-head at back pres-sure p This equation was recently found to contain an error due to an incor-rect expression for variation of vol-ume fraction of gas with pressure. Values calculated by the original equation are close to the correct values for small percentages of gas for which it was derived but become increasingly erroneous as the gas content increases and are completely unusable at the very high gas con-tents employed in aerated mud. DERIVATION The nomenclature used is the same as in Strong&apos;s paper. The pres-sure dP exerted by a lamina of thickness dh at a depth h ft is ex-pressed by Eq. 1 where x = volume fraction of gas at depth h dP = D (l - x)dh . . . (1) When we express x in terms of the wellhead per cent gas n, the well-head pressure p and the pressure P at depth h, the result is Eq. 2. x = (np/p+p)/(100-n) + (np/P + p) When this is substituted in Eq. 1 and rearranged to separate the var-iables the result is dh = 1/D [dp + np/100-n dp/(P + p]. Integration between appropriate lim-its results in or if the wellhead is open so that p = 1 atm and we rearrange Eq. 4, the result is loss in head = hD — P = n/100-n In (P + 1) ...(5) Eq. 5 can be solved easily by successive approximation using hD as a first approximation for P on the right-hand side and calculating a first value for loss in head. This can then be used to get a new value of P using the left-hand side of the equation. Values converge very rap-idly even for quite large values of per cent by volume gas n. EXAMPLES The effect on bottom-hole pressure of 50 per cent gas cutting of the mud in a 5,000-ft well using 10 Ib/ gal mud will be calculated, we have: h = 5,000 ft 10 X 7.48 144 X 14.7 = 0.0353 atm/ft hD = 176.5 atm n = 50 Substituting these values in Eq. 5 and using hD as a first approxima-tion of P on the right, we have, approximate loss in head = 50/50 In (176.5 + 1) = 2.30 log 177.5 = 5.17 atm. We calculate a new value of P as follows:

Jan 1, 1958

By J. M. Sands

Describes important developments in, four counties, all of which brought in 40° oil. Indications are favorable for the future, although the daily production of the agate decreased 19,000 bbl. during the year 1923. THE state of Kansas had a daily average crude-oil production of about 87,000 bbl. at the beginning of the year. This showed a steady but consistent decline during the 12 months, so that the daily average production at the end of the year was slightly over 68,000 barrels. GREENWOOD COUNTY THE most active development in Kansas during the year was in Green-. wood County, in a line from near Sallyards in T 25-R 9 to east of Madison in T 22-R, 11. The main part of. this development has been in a succession of pools beginning at the south with the Polhamus pool in 34-24-9 and 3 and 4-25-9; Thrall pool in 28, 29, 32 and 33-23-10; Burkett pool in. 13, 23, 24 and 26-23-10; Seeley and Wick pool in 22, 27, 28, 32 and 33-22-11 and 5, 6, 7, and 8-23-11; Frohm pool in 11, 14 and 15-22-11; Madison pool in 31 and 32-21-12 and 6, 22, 12 and 1-22-11. Some of these pools, have two or more sands, locally called Bartlesville and Burgess Sands. They range in. depth from 1850 ft. near Madison to 2450 ft. near Sallyards, the general dip being to the west. The thickness of these sands runs from 20 to 100 ft., the initial production of the wells being 50 to 1000 bbl. of 40-gravity oil or better.

Jan 3, 1924

By AIME AIME

THE statements quoted below range widely over the field of contact between education and industry. &apos;Their sources are as indicated. True Education "Education must escape from its traditional warped and narrow views as to what constitutes life and preparation for life. It must train young men and women for vigorous, wholesome development of every important element of personality. It must build physical and mental health. For each student it must inform and discipline thinking in literature, in physical, bio- logical and social sciences, and in philosophy. It must awaken appreciation of beauty.

Jan 1, 1929

By OTTO HERRES

TO operate the bituminous coal industry in the United States in 1929 cost $770,237,000, of which $30,739,000 was paid for purchased power and $34,947,000 for new machinery and equipment. Equipment aggregated 3,124,000 hp. A mining industry that spends $65,000,000 a year for power and machinery is likely to be fairly well mechanized, and when the machinery in use takes three million horsepower to drive it, the probabilities are that mechanization has been in progress in at least some of the mines for quite a few years. Mechanization in its broadest sense implies machine cutting, power drilling, mechanical loading, electric haulage, mechanical sizing and cleaning, and lesser applications of machinery to mining operations. But the introduction and development of mechanical loading has been the feature of coal mining recently and its extraordinary progress is causing great changes within the industry. First, let us look at the records. The quantity of coal loaded mechanically in bituminous mines was:

Jan 1, 1933

Edwin H. MessiteR, New Pork City (communication to the Secretary*):—Under the heading " Flues," Mr. Edwards refers to the Bee-hive construction, a cross-section of which is shown in Fig. 4 of his paper. A flue similar to this was designed by me about six years ago,&apos; in which the walls though much thinner than those described by Mr. Edwards gave entire satisfaction. These walls, from 2.25 in. thick throughout in the smaller flues, to 3.25 in. in the larger, were built by plastering the cement mortar on expanded-metal lath without the use of any forms, or cribs, whatever, at a cost of labor generally less then $1 per sq. yd. of wall. Of course, where plasterers cannot be obtained on reasonable terms, the cement can be molded between wooden forms, though it is difficult to see how it can be done with an interior core only as stated by Mr. Edwards. In regard to the effect of sulphur dioxide and furnace-gases on the cement, I have found that, in certain cases, this is a matter which must be given very careful attention. Where there is sufficient heat to prevent the existence of condensed moisture inside of the flue, there is apparently no action whatever on the cement, but if the concrete is wet, it is rapidly affected by these gases. At points near the furnaces there is generally sufficient heat not only to prevent internal condensation of the aqueous vapor always present in the gases, but also to evaporate water from rain or snow falling on the outside of the flue. Further along, a point is reached where rain-water will percolate through minute cracks caused by expansion and contraction, and reach the interior, even though internal condensation does not occur there in dry weather. From this point to the end of the flue the roof must be coated on the outside with asphalt paint or other impervious material. In very long flues a point may be reached where moisture will condense on the inside of the malls in cold weather. From this point to

Jan 1, 1905

By W. H. Bassett

W H. BASSETT was elected first vice-president at . the executive session of the new Board on Tues- day afternoon. Karl Eilers, H. Foster Bain, Thomas T. Read, and H. A. Maloney were respectively re-elected as treasurer, secretary, assistant secretary and assistant treasurer. The new Executive Committee will consist of W. H. Bassett, Chairman, and Richard Peters, Jr., H. C. Bellinger, E. De Golyer, Edgar Rickard. The Finance Committee will be J. V. W. Reynders, Chair- man, Henry Krumb and H. G. Moulton. George D. Barron was reappointed Chairman of the Admissions Committee, and F. Julius Fohs put in charge of the Papers and Publications Committee.

Jan 1, 1929

By A. A. Bauer, M. S. Farkas, F. A. Rough

An investigation of the d-phase relationships between the uranium-zirconium and uranium-molybdenum systems was conducted. A ternary cut joining U-31.5 at. pct Mo to U-74 at. pct Zr is presented on the basis of differential thermal analysis, metallographic and X-ray diffraction data. HyDothetical isothermal ternary sections, inferred from the determined d-phase relationships, are presented for the uraniztm-zirconium-molybdenum system. A ternary eutectoid, at 454ºC, was found in zirconium-rich alloys. Its composition, based on metallography and electron-beam micro-analysis, is estimated as U-61.7 at. pct Zr-7.1 at. pct Mo. URANIUM binary systems have been studied extensively in the past to supply basic information needed in reactor-fuel development. Alloys of uranium with zirconium, molybdenum, and titanium are of particular interest because, in these binaries, extensive or complete solid solubility between the body-centered-cubic forms of these elements and y uranium exists. In addition, an intermediate phase of limited solid solubility forms in each of these systems by transformation of the high-temperature ?-uranium phase. Previous investigation of the phase relationships existing between the intermediate 6 phases in the uranium-zirconium-titanium and the uranium-molyb- denum-titanium ternaries have already been reported.1,2 The results of an investigation of the phase relationships existing between the 6 phases in the uranium-zirconium-molybdenum system is reported here. The phase equilibria that occur in the solid-state portions of the pertinent binary systems are shown in Fig. 1. Phase nomenclature appearing throughout the report is as shown in these diagrams. The crystal structure of the phases in this system are given in Table I. EXPERIMENTAL PROCEDURES Alloys were cast and fabricated to 1/4-in.-diam rod. Specimens were heat treated and examined by metallographic and X-ray techniques. Thermal-analysis data were obtained.

Jan 1, 1960

By W. M. Bertholf

To establish the minimum or break-even efficiency of a preparation plant it is necessary to determine the ratio of production costs per ton of run-of-mine coal to sale price per ton of clean coal. This ratio is then divided by the percentage of plant recovery. Assume, for example, that washability data indicate that 80 pct of the run-of-mine coal may be recovered as clean coal selling for $5.50 per ton. If the same plant incurs $4.00 per ton mining costs and $0.25 per ton preparation costs, the cost to sale price ratio would be 4.25/5.50, or 77.3 pct. Since the plant is capable of recovering 80 pct of the run-of-mine coal the minimum plant efficiency would be 77.3/80, or 96.6 pct. If the plant efficiency were increased to 100 pct the limit on preparation costs would be increased to $0.40. Therefore, the washing operation must be accomplished at a cost of $0.25 to $0.40 per ton and the efficiency must be 96.6 pct or higher. By plotting a curve of efficiency vs preparation costs and taking into consideration the plant cost, feed rate, operating costs, and yields of the various flowsheets it is possible to select a financially sound process. Table I classifies preparation processes according to their nature of separation. It must be remembered that what is thought to be ideal does not always prove practical under actual operation. An outstanding example is the work done on the plunger-type jig. It was determined by the law of equivalence that this unit could treat satisfactorily only sizes differing by no more than a ratio of 2. Yet these same plunger-type jigs are successfully treating sizes considerably greater than this ratio.

Jan 12, 1957

By F. B. Plummer

PROGRESS during 1940 in oil-production technology has been confined largely to a steady advancement in practices inaugurated in previous years, rather than the introduction of any new startling process or technique. The usual changes, improvements and additions in production equipment have been introduced. An increased attention to, and interest in, research has been noteworthy, resulting in improved efficiency in operation of oil wells. control, utilization of reservoir energies and reduction of production costs.

Jan 1, 1941

By H. K. Najarian

AT Josephtown, Pa., on the south hank of the Ohio River about thirty miles northwest of Pittsburgh, is the Josephtown electrothermic zinc smelter, which operates as a custom smelter receiving both domestic and foreign concentrates. Constructed in 1930, the plant was designed to produce zinc oxide and to process 124 tons per day of zinc concentrate obtained mainly from the Company&apos;s New York State mines. Zinc condensing equipment was developed in conjunction with the St. Joe electrothermic. furnace and commercial production of zinc metal started in 1936. Expansion of the plant and equipment during 1939 and 1940 brought the processing. capacity to upwards of 300 tons of concentrate per day; and the plant operated at peak capacity during the war years, earning the Army-Navy "E" award and three subsequent stars. A postwar plant expansion is now under way, designed to increase the capacity to 500 tons per day of concentrate, with an estimated daily production of 160 tons of zinc metal of various grades, about 100 tons of zinc oxide, and 425 tons of sulphuric acid as major products.

Jan 1, 1947

By E. J. Wasp, N. T. Cowper, R. A. Davis, W. F. McDermott

On October 26, 1967, the world&apos;s first long distance iron ore slurry line was put into service by Savage River Mines. The commissioning of this $5 million installation climaxed three years of development, testing, and construction. The mine and concentrator are located approximately 30 miles from the town of Waratah in a rugged semi-mountainous terrain. The open pit mine will produce approximately 5 million long tons a year of crude ore, which is ground in autogenous mills, separated magnetically, thickened and prepared for the pipeline. The 53-mile pipeline transports the 2,250,000 tons of concentrate north to Port Latta. The Port Latta facilities include the receiving, storage and thickening equipment for the slurry as well as filtering, balling and pelletizing equipment to produce the 2,250,000 tons of high grade iron ore pellets. Five vertical shaft furnaces are used for pelletizing.

Jan 1, 1969

By J. T. Ryan

FIGURES for the first 11 months of 1935 indicate that the total coal production of the United States for 1935 will be approximately 416,000,000 tons, or almost identical with the production figures for 1934. Bituminous production will show a gain of about six million tons, and anthracite production a loss of about the same tonnage, compared with the previous year. This does not take into account so-called "bootleg" anthracite coal amounting to approximately 5,000,000 tons. Twenty-nine years ago, in 1906, the production of coal was almost identical with this year&apos;s production. At that time, 626,000 men were employed in producing this tonnage, which took a toll of 2138 lives from accidents. In 1935, approximately 540,000 men were required to produce the same tonnage with an estimated loss of 1155 lives. This shows definite safety advancement in the industry, but even so accidents are still costing an average of 3c. a ton more than they should.

Jan 1, 1936