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By Charles E. Mann, Douglas J. Lootens, Rufino B. Bomasang

INTRODUCTION The economy of the Philippines, like that of many other developing nations, has been severely strained by the rapid escalation of imported crude oil prices over the past decade. In an effort to reduce its dependence on imported crude, significantly increased emphasis, both in the public and private sectors, is being placed on the rapid development of Philippine coal resources. This paper is designed to present an overview of the Philippine coal industry, its historical development, present status, and future potential. COAL RESOURCES AND RESERVES In 1977, a British consulting firm collated all available data on Philippine coal resources and estimated the total coals potential at about 0.9 to 1.4 billion metric tons (mt). In 1981, using more recent data the Philippine Bureau of Energy Development (BED) estimated the potential at 1.7 billion mt. The BED estimated the proven coal reserves at 257 million mt. These coal resources are scattered throughout the major islands comprising the Philippine archipelago. However, about 52 percent of the currently proven reserves are located in the tiny island of Semirara between the islands of Mindoro and Panay. About 25 percent is located in the Cagayan Valley in northeastern Luzon while 13 percent is located in Mindanao. The remaining 10 percent is scattered in the islands of Cebu, Samar, Mindoro, Negros, Polillo, Batan, and Catanduanes. The bulk of Philippine coal is sub-bituminous in rank. Some bituminous coal with coking properties has been reported in Malangas, Zamboanga del Sur in western Mindanao, and in Catanduanes, but the exact tonnages have not been accurately established. It is believed, however, to be at most 5 million tons. A little semi-anthracite coal exists in Malangas with the remaining unmined reserves estimated at a little more than one million tons. Table 1 shows the quality ranges of Philippine coal deposits.. Figure 1 shows the distribution of known Philippine coal resources PHILIPPINE COAL PRODUCTION In 1981, total coal production was 330,720 mt. Though modest by world standards, this was the highest production figure in Philippine coal mining history. This came from 25 mines, mainly small scale underground mines of capacities varying from 10 to 100 tons per day (tpd). The actual production capability was about 800,000 mt in 1981, but the producers had to scale down production to the level of local demand. Even then, some 50,000 tons of coal remained unused by year-end. The 1981 production came from the following areas: Cebu 224,751 mt (68%) Zamboanga del Sur 57,095 mt (17%) Semirara 13,222 mt (4%) Polillo 17,132 mt (5%) Cebu has always been the main source of Philippine coal production. Coal mining has been carried out in Cebu for the last 50 years, although it has been characterized by a very crude mining technology. Due to the limited size of coal deposits and complicated geology, however, Cebu's share of the total national production is expected to go down to only 10 percent by 1985. Except for Semirara's open-pit production, the balance of Philippine coal production comes from underground mines. These underground mines generally use room-and-pillar methods, except for those at Zamboanga del Sur and Polillo which use modified longwall methods patterned after those used in Taiwan. Due to the much higher mining recovery, most underground mines are now planning on switching to longwall. At Semirara, an open pit mine is now being developed for a one million ton per year capacity by 1984, and will then be the largest Philippine coal mine. While present earthmoving work is done with bulldozers, loaders, and trucks, the main coal extraction equipment will be bucket wheel excavators imported from Austria. PRESENT COAL UTILIZATION Existing coal production is sold mainly to coal- fired cement plants and power plants. The balance is sold to a sugar central, a coconut oil mill, a food processing plant, and two plants producing carbide and ferro-silicon alloys.

Jan 1, 1982

By S. R. Mitchell, M. C. Fuerstenau, A. M. Gaudin

The two principal attributes of a thickener pulp are its settling rate and the ultimate pulp density of the thickened mud. Testing for evaluation of thickening attributes of a pulp has usually been done batch-wise by placing the pulp in a one-liter glass graduated cylinder and noting the position of the boundary between clear liquor and mud as a function of time. Refinements in testing have included repulping a thickened mud to various dilutions and evaluating the settling rate on these various repulped muds. This is considered to make possible an estimate of the settling rate as a function of pulp density and therefore to pinpoint the slowest settling pulp which ought to be critical from the standpoint of thickener plant design. Finally, tests are sometimes made with model-size continuous thickeners of arbitrary size. In batch tests it is generally accepted that the behavior of a pulp is described by a curve such as the one shown in Fig. 1, in which the position of the boundary between clear liquor and mud (the mudline) is plotted against time. It has also been assumed that the initial settling rate (the slope of the line segment AB) is descriptive of the pulp and independent of the size and shape of the container. Where the settling rate is slower and clearly a function of time (portion CD of the settling curve), it has been assumed that the data are not of much use because of the intrusion of compacting phenomena. The next important datum is the height. PE, which is a measure of the so-called ultimate pulp density. This quantity is assumed to be characteristic of the pulp and independent of the container. When actual data are recorded, results are rarely exactly according to this theory and numerous deviations may be noted. For example, some pulps seem to require an appreciable delay before beginning to settle, as in line AF. This is spoken of as representing the need for an induction time for flocculation. It has seemed desirable to explore thickener data with a more sensitive charting tool than is afforded by the coordinates of Fig. 1. Instead of relating the position of the mudline to time, it has seemed desirable to relate the rate of descent of the mudline to time. This, in effect, amounts to charting the slope of the line in Fig. 1, or the derivative of its ordinates, against the same abscissas. Fig. 2 shows the new chart for the curve ABCDE of Fig. 1. When this new tool was applied to data obtained, for example, with sized clay pulps, astonishing differences appeared. Apparently the traditional tool conceals differences instead of revealing them, whereas the new method emphasizes them. It seems, also, that both charting methods have their uses. In this article attention is focused on the effects of pulp density and container height as they affect the settling behavior of flocculated suspensions of a sized kaolin. The thesis by M. C. Fuerstenau' gives more details than this summary, as well as some information about the effect of other variables. Sized fractions of kaolin were used, supplied through the courtesy of S. C. Lyons of the Georgia Kaolin Co. Large batches were made by agitation in a flotation machine and in a glass jar equipped with a mixer. The settling vessels were 2, 4, and 6-in. diam glass cylinders of various heights. Pulp Dilution: The effects of pulp dilution on sedimentation behavior are shown in Figs. 3, 4, and 5. Fig. 3 shows data on pulps ranging from 2.44 to 16.7 pct solids initial pulp density, flocculated with Separan and charted in the traditional manner. At least two of the lines appear nearly straight in their early sections, but neither of the others does. Each of the curves in this figure includes a region labeled

Jan 1, 1960

By Carville E. Sparks, Rollin Farmin

TRIAL installations of rock bolts, of the slit-rod-and-wedge type, were under way at several units of Day Mines, Inc., when Korean hostilities interrupted the already slow deliveries of steel bars to the Coeur d'Alene district. Factory-made bolts had not yet been put on the local market, so the program was halted for lack of supplies. Interest was revived by a visitor's description of wooden roof bolts. These were said to have been used briefly with apparent success in a coal mine, until apprehension voiced by the U. S. Bureau of Mines caused the practice to be suspended. To make wooden bolts for trial in ground support, Day Mines acquired a second-hand doweling machine equipped with two cutting heads, one to turn out the desired round rods of 2-in. diam, the other to turn out 1-in. rods to be used as powder-tamping sticks. This machine was installed in the all-weather sawmill of the Hercules mine unit at Burke, Idaho, where fabrication of the wooden bolts commenced early in 1951. Most of the mining in the Coeur d'Alene district is along steeply dipping veins in shaly quartzite and argillite of Algonkian age. Ground support commonly is required in zones where the rocks have been sheared, brecciated, and hydrothermally altered. Pressure from the sidewalls is more troublesome than weight overhead, but both increase with the size of the mine opening. Caving may come from a progressive sloughing of irregular rock fragments or from an exfoliation and buckling of the layered wall rocks. The disintegration is thought to develop from an initial elastic expansion of the rock toward the newly-created mine opening, followed by the dilation of many tiny partings in the rock by absorption of water. As the partings widen, masses of rock develop weight and become free to fall. The function of rock bolts is to prevent or retard widening of partings in the rock supported. Wooden Bolts, Wedges and Headboards Bolt assembly used by Day Mines consists of a bolt 4 or 6 ft long, two wedges 16 in. long, and a headboard 30 in. long, Fig. 1. All four pieces are made of local red (Douglas) fir, either green or well-soaked in the mill pond before it enters the sawmill. Bolts are fabricated from cants, 2 1/4 in. sq, cut from relatively straight-grained timber with a minimum of knots and trimmed to 4- and 6-ft lengths. The bolt then is turned in the doweling machine from 21/4 in. sq to 2 in. diam round, except for a 4-in. length at one end which is left full square to provide the striking head and the shoulder that holds the headboard in position for wedging. The foot end of the bolt is slit with a thin saw for a length of about 16 in., thereby making a slot to receive the wedge against which the bolt is driven for anchorage at the bottom of the rock hole. A similar slit, 12 in. long, is made in the opposite (head) end of the bolt to receive the second wedge, which crowds the headboard against the ground at the collar of the rock hole and puts the bolt in tension. The second slot is aligned 90" from the plane of the first slot to avoid Longitudinal splitting and is notched out slightly to allow easier insertion of the collar wedge after the bolt has been driven to bottom. To prevent splitting the headboard by spreading action of the head wedge, this slot is oriented at 90" to the grain of the headboard when the pieces are assembled, Fig. 2. The wedges are similar to standard mine wedges, but more slender; they are cut 1 7/8 in. wide and 1 in. thick at the heel and taper out in 16 in. of length. The headboard, or bearing plate, is not necessary for some types of ground but generally is desirable because it helps the bolt to support an area of loose, friable rock and reduces the tendency for the rock at the collar of the hole to split away from the wedged head by distributing the pressure over a wider rock surface. The headboard may be a 24-to 30-in. length of 3-in. plank, 8 to 12 in. wide, but a similar length of rounded sawmill slab serves equally well at 20 pct of the cost. A hole of 2-in. diam is bored or punched through the center of the headboard, either at 90" or at various high angles to its surface. The bolt is inserted to its shoulder through this hole, then driven into the rock hole. Bolts, wedges, and headboards are given a full timber preservative treatment to inhibit rot. Bundles of each are immersed in a warm saturated solution of Osmose salts in water for 48 hr, removed, dripped dry, and stored in a relatively humid underground depot to cure. Most wooden rock bolts used by Day Mines are 4 ft long. Holes to receive them, about 42 in deep and 2 1/8 in. in diam, are drilled into the rock' to be supported, nearly normal to the periphery of the mine opening. The type of drill used is dictated by convenience: stoper, jackleg, or jumbo-mounted drifter. Correct depth of the hole is assured by use of a measuring stick that has been cut to the proper distance from drill chuck to the ground at the collar of the hole when a standard length drill rod is at the bottom. The bolt is seated to the shoulder through the hole in the headboard, the foot-end wedge is placed in its slot, and the assembly is inserted into the rock hole. Then the bolt is driven until it is seated solidly on the wedge against the bottom of the rock hole. Driving may be by hand with a sledge, or

Jan 1, 1954

By Murray F. Hawkins, W. J. Ainsworth

In all types of subsurface pressure gauges the extension which occurs in the pressure-sensitive element is a function of the difference between the external (well or calibration) pressure and the internal pressure within the gauge, rather than a function of the external pressure only. The internal pressure is near atmospheric and depends upon (a) the quantity of air sealed within the gauge at the time of calibration or measurement, (b) the quantity of moisture (liquid water), if any, sealed within the gauge, and (c) the temperature at which the calibration or well measurement is made. Part of this correction for the change of internal pressure with temperature is taken care of by the customary temperature coefficient of the gauge. However, part of it is not, and while this portion may be only a few psi, it is nevertheless predictable or preventable, and should be considered in precision measurements. ERROR NO. 1 If air is sealed in the gauge at the same temperature and pressure for both the calibration and the well measurements, the usual temperature correction will take care of any difference between calibration and well measurement temperatures. However, if air is sealed within the gauge at temperature T1 and pressure P1 at calibration but at temperature T2 and pressure P2 for a well measurement, because different amounts of air are sealed within the gauge in each case, the internal pressure at. or corrected to, calibration temperature Tr will he different by where all temperatures and pressures are absolute. The calibration temperature is used, and not the well measurement temperature, because the usual temperature correction reduces the well measurements to calibration temperature. The correction term as calculated by the above equation is separate from. and in addition to, the usual temperature correction. Example: T, = 540°R, sealing temperature at calibration P, = 14.7 psia, sealing pressure at calibration T2 = 460°R, sealing temperature at well P2 = 14.7 psia, sealing pressure at well Tr = 660°R, calibration temperature AP = 660 [14.7/460 — 14.7/540] = 3.1 psi While this error is small even under these somewhat maximal conditions, it nevertheless represents a practical situation which did occur, and which as a matter of fact gave rise to this note. Where AP is positive, as above, the correction is added to the measured pressure; where negative, subtracted from the measured pressure. This correction should also be considered in successive calibration runs where the gauge, for example, may be warm from a previous calibration at an elevated temperature. ERROR NO. 2 Where a small quantity of moisture (liquid water) is sealed within the gauge at atmospheric conditions, the increased vapor pressure of the water at higher well or calibration temperatures will cause an increase in internal pressure. This moisture will come presumably from condensation within the gauge following temperature changes, from moisture on the operator's hands, and from atmospheric moisture (rain, mist, fog, etc.). Calculation shows that approximately 0.2 cc of water (three to four drops) is sufficient to saturate the air within an Amerada RPG-3 Gauge at 160°F, at which temperature the vapor pressure of water is about 5 psia. As the vaporization occurs in a sealed volume, the increase in internal pressure will be in excess of this 5 psi. At higher temperatures the pressures will be higher; however more water will be required to saturate the air within the gauge. Some experimental work was carried out with an Amerada RPG-3 Gauge at 200°F fitted with a 1,000 psi element, both with a dry recording chamber and with a small amount of water added. The results directly proved the existence of the error due to the presence of moisture, and, it is felt, indirectly, due to the differences in sealing temperatures and pressures, as both effects may be ascribed simply to an increase in the moles of gas within the recording chamber. SUMMARY In precision measurements the error introduced by sealing the gauge during a well test at a different temperature and pressure from that of calibration may be corrected for by using the equation presented, or it may be prevented by taking care always to seal the gauge at near calibration conditions. The error introduced by sealing moisture in the gauge may be prevented by taking care to keep moisture out of the gauge, or by removing the moisture by either warming or evacuating the gauge. Both of these errors are independent of the range of pressure measurement and the type of gauge, and are in addition to the usual temperature correction. ACKNOWLEDGMENT Appreciation is expressed to W. B. Kendall, Geophysical Research Corp., Tulsa, Okla., who pointed out the error from differences in sealing temperatures during some winter work in Canada. ***

Jan 1, 1956

By W. H. Bassett

IN the commercial annealing of cartridge brass there are four points regarding which definite data are essential. They have to do with the correct interpretation of grain count in its relation to annealing temperature and, incidentally, to Brinell hardness. These points are: 1. The comparison of the grain sizes of two cartridge-brass mixtures: 69 copper, 31 zinc, 0.376-in. (9.5-min.) gage; and 68 copper, 32 zinc, 0.130-in. (3.3-mm.) gage. 2. The comparison of the grain sizes of annealed metal that had previously been reduced by rolling varying amounts; for instance, 20.0, 36.6, 50.9, and 59.1 per cent. 3. The determination of standards for grain sizes on annealed brass of the composition 68 per cent. copper, 32 per cent. zinc and 69 copper, 31 zinc. 4. The comparison of grain size with Brinell hardness on identical samples of annealed metal. In their comprehensive and thorough investigation of the recrystallization of cold-worked alpha brass on annealing,1 Mathewson and Phillips have discussed the relations between temperature of anneal, degree of deformation, and structural alteration in alpha brass. They have also shown certain comparisons between the ordinary physical properties and the grain size of annealed brass. The purpose of the present investigation is mainly concerned with the grain size of cartridge brass, its relation to Brinell hardness, and the publication of sufficient data to enable those engaged in the inspection of such material to have a correct foundation upon which to work.

Jan 1, 1919

By B. S. Norris, Charles R. Austin

IN a recent paper the authors1 discussed the reactions to tempering of hypereutectoid steels quenched from 1000° C., as revealed by studies on changes in hardness, electrical resistivity, coercive force and metallographic characteristics. The data, which related specifically to tempering temperatures of 550°, 650° and 710° C., for varying periods up to 125 hr., permitted the following important conclusions: 1. The rate of softening of the steels at the temperatures examined differed markedly although the steels were of "similar" general chemical composition. 2. The temperature of maximum softening spread over a range from about 650° C. to the eutectoid inversion. 3. The divergence in behavior was ascribed, at least in part, to a profound difference in tendency to graphitization on annealing at subeutectoid temperatures. 4. The node of spheroidization of the cementite, as revealed by metallographic studies, also varied considerably in the different steels. The purpose of the present investigation is to pursue further the study of the difference in behavior of the steels on tempering over a wider range of temperature. To this end the following features have received particular attention: 1. Comparison of temperatures of maximum softening of the various steels when the quenched samples were tempered for 5 hr. and for 75 hr. followed by furnace cooling and by quenching in water. 2. Determination of the effect of quenching temperature, prior to tempering, on the rate of softening. Quenching temperatures of 1000° and 850° C. were used in the investigations. 3. Examination of the microstructures obtained with the various heat-treatments employed.

Jan 1, 1938

By S. E. Bronisz, J. O. Brittain

Experiments have been carried out to study the effect of an applied stress on strain aging in steel. It was found that the stress-induced strain aging reaction proceeds rapidly in the temperature range of 0° to 60°C. The variation of stress-induced strain aging with plastic strain, aging time, and temperatures is reported. The activation energy for the stress-induced strain aging was found to be 10,000 cal per mol for the steel used in this investigation. VARIOUS observations1-5 have been made of the influence of an applied stress upon the yielding of iron and steel. However, there have been only cursory studies of the effect of an applied tensile stress upon strain aging in iron and steel1,2 The following represents the results of a preliminary study of strain aging as influenced by the application of an external tensile load during the aging process. For the most part, previous experimental work on strain aging can be classified into two parts. First, the experimental technique involved a direct determination of mechanical properties of a suitable specimen that has been plastically deformed to a constant strain and aged generally at a temperature of the order of 200°C, with the aging times varying from about 1/10 to 2 hr.6-8 The second classification of strain aging experiments involves an indirect determination of strain aging by means of physical property such as electrical resistivity or internal friction as a function of the aging time at a constant temperature and strain.9-12 The latter results have

Jan 1, 1961

By S. Nenno, T. Saburi, Y. Mizutani, M. Yomarnoto, K. Komatsu

In the equilibrium phase diagram of the Ni-Mo sys-tem1,2 available at present, intermetallic compounds (or ordered phases)ß(Ni4Mo) and ?(Ni3Mo) are known to exist to the nickel rich side. In our recent investigations on Ni-Mo alloys, however, it has been found that the quenched alloys from the high temperature region containing more molybdenum than the Ni4Mo do not decompose directly into 0 and when annealed isothermally below the peritectoid temperature, i.e., 865°C. This note is to report the existence of Pt2Mo type superlattice (possibly ordered Ni2Mo) as a metastable phase in the early stage of decomposition. Alloys containing 20.1, 22.7, and 24.4 at. pct Mo were prepared by vacuum-melting electrolytic nickel of 99.9 pct purity and molybdenum powder of 99.9 pct purity in an alumina crucible. From the resulting ingots, sheets 0.3 mm thick were obtained by forging, hot-, and cold-rolling. The sheet specimens were water-quenched from ! region and subsequently annealed for various periods of time at temperatures below the peritectoid temperature. On these specimens were made optical-microscopic and transmis-sion-electron-microscopic observations. Fig. 1- Optical micrograph of the 22.7 at. pct Mo specimen annealed for 100 hr at 700°C after quenching. (ß + ?) region, B, nucleated along a grain boundaries are invading the preexisting (ß + Ni2Mo) region, A. Details of A are observable only by electron microscopy. Etchant: 60 pct HNO3-40 pct CH3COOH.

Jan 1, 1970

By Charles E. Mauk, H. William Prengle, Robert W. Legan

A process breakthrough was needed for the complete destruction of total cyanide influent as high as 100,000 mg per 1, or below 1 mg per 1, to produce an influent with total cyanide below the limit of detection. The process uses ozone for the total destruction of cyanide and metal cyanide complexes. It was found that while ozone alone reacted with the iron cyanide complexes too slowly for practical application, ozone at slightly elevated temperature became more effective, and ozone with ultraviolet light became effective enough to permit the design of a successful system.

Jan 1, 1977

By M. Dubrovsky, J. W. Evans

A laboratory scale cell (approximately 50 amps current) equipped with a fluidized bed cathode was used in a study of the electrodeposition of cobalt from acidified aqueous sulfate solutions The effect of current density (in the range 0.1 to 1 amp/cm2 of diaphragm), cobalt concentration (4 8 to 100 glliter) and catholyte pH (1 35 to 6 2) on current efficiency and energy consumption were determined. Generally the current efficiency increased with increasing current density and cobalt concentration Current efficiencies were comparable to those reported in the literature for conventional electrowinning

Jan 1, 1981

By William E. Ford, Edward Salisbury Dana

The Nitrates being largely soluble in water play but an unimportant r81e i~ Mineralogy. SODA NITER. Rhombohedral. Axis c = 0.8276; rr" 1071 A 7101 = 73' 30'. Homeo- morphous with calcite. Usually massive form, as an incrustation or in beds. Cleavage: r (1011) perfect. Fracture conchoidal, seldom observable. Rather sectile. H. = 1.5-2. G. = 2'262.29. Luster vitreous. Color white; also reddish brown, gray and lemon-yellow. Transparent. Taste cooling. Optically -. w = 1'5874, e = 1'3361. Camp. - Sodium nitrate, NaN03 = Nitrogen pentoxide 63'5, soda 36'5 = 100.

Jan 1, 1922

By Lawrence Adler

Any design of a mine roof in bedded deposits which ignores differential deflections at the supports can quickly lead to dangerous overstressing. As illustrated by the typical case presented on page 1029, if one support deflects only 0.1 in. more than its neighbor, stresses 50 pct higher than anticipated will result. Previous studies1-3 have assumed that the mine roof acts like a beam clamped at both ends, with no deflection or slope4 at the supports (Fig. 1). Essentially this assumes:5,6 [BA = OD = 0 and dA = dB = 0.]

Jan 10, 1959

By J. E. Chappelear

We are concerned with the problem of calculating the forces produced by passing waves on a submerged object (in particular a drilling barge) located on the ocean floor. We employ the linearized theory of waves (the Airy1 theory), since solutions are generally more easily obtained in such a theory. It will be indicated how an approximate solution to the problem can be obtained for waves of finite height by use of a linear combination of Airy waves. We shall assume the submerged object has an infinite length, thus neglecting diffraction and making the problem two-dimensional. We find that there is not a considerable difference between our solution and the usual engineering solution. The force, as a function of time, is plotted in Figs. 4 and 5, together with the usual solution for easy comparison. So far, no experiments have been found to compare with the results. As will be shown, the calculations can be extended to waves of finite height (Stokes2-Struik3 waves) in an approximate fashion. This can be done by calculating the force produced by each Fourier component of the wave, separately, and omitting the coupling between the various harmonics produced by Bernoulli's theorem. The problem and its boundary conditions will be briefly stated. We seek a function, +, the velocity potential, which is harmonic, ?2?=0...........(1) and which satisfies the boundary conditions ??/?n =0 .......(2) (the normal derivative of + is zero) on the ocean floor and the sides of the barge, and s??/?y = ?..........(3) on the ocean surface, where s = g/w2........(4) The physical reasons for these choices and an interpretation of ? are given by Lamb.' As previously stated, we assume also that the problem is two-dimensional. We set up our coordinates such that x = 0 is the center of the barge and y = 0 is the ocean floor. The coordinates are shown in Fig. 1, together with the location of the barge on the bottom. Our solution may be decomposed into a function which is symmetric about x = 0,?e, and one which is antisymmetric, +,, ? = ? + ?o...........(5) Thus directly above the barge we have as boundary conditions that for x = 0, ?o = 0, ...........(6) and ??e/?x = 0........(7) Now, far from the barge, we have that (for x positive and assuming the waves are incident from the left) ?e ? Ae cosh ky cos k (x + a) , . . . (8) and ?e ? Ao cosh ky cos k (x + ß) . . . . (9) where + means ". .. approaches ... for x large." Let us put Ae, Ao, and k equal to one, and remember that, in accordance with the derivation of Eq. 3, a sinusoidal time factor is to be supplied. Then we may rewrite Eq. 8 for ?e, and we find that

Jan 1, 1958

By WOOLSEY MCA JOHNSON

INDEPENDENTLY of the recognized chemical reactions involved in the production of metallic zinc, the process is affected by physical conditions in efficiency, and by commercial as well as technical economy. To offer some observations concerning these conditions is the purpose of the present paper. Among the important elements of this problem is the physical nature of the particles constituting the mass of zinc oxide to be reduced. This may be the result, either of the manner of the original deposition of the mineral, or of the treatment to which it has been subjected in the roasting-furnace. The physical character of the reducing-agent is another important factor, since the carbon of some coals is much more active in reduction than that of others. The degree of fineness, whether of ore or coal, or both, likewise plays a significant part. These factors, together with the physical admixture of foreign substances with the ore, may give rise also to chemical "side-reactions," sometimes hindering or retarding, and sometimes accelerating, the reduction. Finally, thermal conditions, including the relative conductivity of the materials, deserve to be taken into account. In firing a furnace with retorts charged with a mixture of roasted ore and coal, the first reaction is naturally the boiling-off of the contained water ; the next is the distillation of the light hydrocarbons from the coal; the next is the reduction of the iron oxide to protoxide and sub-oxide successively, and then to metallic sponge; and the final reaction is the reduction of the zinc oxide in the ore by the carbon to metallic zinc. This reaction commences at 1,022° to 1,060° C., according to the physical nature and condition of the coal and ore. Since the retort and the charge are far from being good con-

Sep 1, 1907

By L. W. Bahney

FREQUENTLY in using a thermo-electric pyrometer for measuring the temperature of a furnace, a hole is drilled at the back or side of the furnace, through which is introduced the tube containing the thermo-couple. At times the couple is left almost where it drops, for the reason that it soon becomes too hot to be handled easily, and the space at the back or side of the furnace may be so small and uncomfortably hot that an easy and accurate adjustment is nearly impossible. In order to overcome this clumsy and unscientific method of using the thermo-electric pyrometer, I designed an adjustable stand, to hold the clay or quartz tube unclosing the platinum and plati-num-rhodium leads that constitute the couple of the Le Chatelier thermo-electric pyrometer. This stand also allows an easy and rapid adjustment of position to varying heights and angles, as may be desired for special reading. This stand is illustrated in Fig. 1. The couple in the tube might be of better service if the tube were held a little higher above the charge in a roasting-dish, changing the height and angle if desired. The stand, made of 1.5-in, water-pipe and mounted on casters, is heavy enough so it cannot be tipped over easily. It is adjusted to any position, and has a universal clamp which holds the clay or quartz protecting-tube. To avoid disturbing the wires of the couple at the cold junction each time the tube is moved, the cooling-bottle sets in a wire basket so placed that it moves forward or back with the couple. To overcome the necessity of leveling and adjusting a galvanometer at each furnace, a switch-board is used to connect

Jan 1, 1910

By W. C. Snow

The plant of the Electrolytic Zinc Company of Australasia Limited, situated at Risdon on the estuary of the River Derwent, near Hobart, was designed and erected primarily to produce high-grade zinc from zinc concentrates produced by the mines of Broken Hill, N. S. W. Consequently the metallurgy of the process has been influenced in its development by the peculiar characteristics of the Broken Hill concentrate, which on treatment gives into solution relatively large amounts of cobalt, manganese and silica. The Risdon works (Fig. 1) produces high-grade electrolytic zinc, leaded electrolytic zinc, zinc dust, cadmium, sulphuric acid and superphosphate. This paper, however, is concerned only with the zinc plant and process. Roasting The roasting of raw concentrate is not a major feature of Risdon's process, and in this respect practice is unlike that of most zinc plants. The roasting is mostly done on the Australian mainland at centers where the sulphuric acid made from the roaster gases can be used for superphosphate manufacture. A small amount, approximately 40 tons per day, is roasted at Risdon, acid being manufactured from the gases. The amount that can be roasted there is limited by the acid requirements of the zinc plant and of the superphosphate plant. Roasting is not carried to completion at the mainland plants, and the product, known as preroast calcine, largely constitutes the raw material for the Risdon zinc plant. It contains about 6 per cent of sulphur as sulphide. Roasting is completed at Risdon for reasons that are set out later in this paper. The zinc plant proper is divided into four main divisions: (1) final roasting division, (2) leaching and purification division, (3) electrolytic division, (4) melting and casting division. There are two subsidiary divisions—the cadmium division and the preroast, acid and superphosphate division.

Jan 1, 1937

By C. E. Sims, C. A. Zapffe

AN investigation of the behavior of inclusions in steel several years ago1 led to the conclusion that some of the commonly occurring inclusions in steel have appreciable solubilities, particularly in liquid steel. In the belief that further study and clarification of this matter was warranted, a laboratory study of oxides in liquid steel was started. These oxides are generally the result of deoxidation treatment, and the basis of an understanding of their behavior must be, therefore, the behavior of the deoxidizer in liquid iron. The Fe-Si-O system was selected for the first study, and the apparatus used and experiments conducted are described herein. APPARATUS The apparatus is shown in Fig. I and the system is shown by diagram in Fig. 2. Furnace Atmosphere Commercial hydrogen (1) passes through an orifice (2) into a manometric system that measures the tankside pressure (3) and the back pressure (4) in the system. A water trap (5) guards against excess tankside pressure. A by-pass with stopcocks (6 and 7) is provided, so that the hydrogen can be routed directly from the tank to the furnace if desired. When stopcock 7 is closed and stopcock 6 is open, the gas proceeds through the conditioning train, of which the first unit is a furnace (8) that contains platinized asbestos at 350°C. The temperature is measured by a thermo-couple and potentiometer (9) and is controlled by a slidewire resistance (10). In [ ] this furnace free oxygen is converted to water vapor. The exiting gas passes over a water bath (11) kept slightly below boiling temperature by an auxiliary heating unit (12). This unit also heats a surrounding copper coil (13) containing water, and the heated water rises through a collar (14)

Jan 1, 1942

By John Dilworth

THIS paper treats primarily of the valuation of developed coal properties by the method of capitalizing their estimated average future earnings. However, in it reference is also made to valuations of undeveloped coal lands and to various methods of appraisal other than by future earnings. FUNDAMENTAL VALUATION PRINCIPLES Before discussing in detail coal property valuations, a few statements bearing on the general subject will be made to present certain of the writer's views on the question as a whole. 1. A certain property at a given time has but, one true value. This is not universally admitted. The appraiser rather frequently finds that a coal owner has in mind three valuations for his holdings; one, the figure at which he feels they should be appraised for taxation purposes; another, the much higher one that should be used if a loan is to he negotiated on the property; and a third, intermediate between the other two, the price for which he would be willing to dispose of his holdings were they for sale. The last is usually nearest to the true value, and because this is so the statement is generally accepted that the fair value of such a property is the price arrived at by negotiation between a willing seller and a willing buyer who are otherwise disinterested. 2. Notwithstanding the foregoing, it is only very generally and approximately true that the proper measure of the value of a given developed1 coal acreage is the unit or per-acre price for which one or more bodies of similar coal in the salve district have been sold recently. To give a few reasons why such a measure of value can not be strictly applied, it may be noted:

Jan 1, 1927

By G. C. Gould

The nature of freckle segregation is determined by chemical analyses, microradiograplzy, and electron microprobe. In addition, the influence of chemistry variation on freckle formation is studied in laboratory experiments. It is demonstrated that controlled laboratory solidification can produce freckle segregation. "FRECKLE" segregation is a discontinuous, rod-like segregation observed in vacuum consumable -electrode ingots. Figs. 1 and 2 show this segregation as pencil-like, dark etching areas of segregation found in longitudinally and transversely sectioned consumable-electrode ingots. In most cases the segregation is perpendicular with respect to gravity and parallel to the longitudinal axis of the ingot. The segregation is most often encountered at or around midradius in large vacuum consumable-electrode ingots. The austenitic alloys, A-286* and 718, † have shown particular susceptibility to this mode of segregation. The origin of these rodlike pockets of segregation is not well-understood. It has been suggested that the segregation is a result of a portion of the electrode dropping off and passing through the arc un-melted. On the other hand, smeltzer1 has demonstrated the freckle segregation in A-286 to be comprised of low-melting constituents relative to the matrix melting point. The discontinuous nature of the segregation is the most perplexing facet of the freckle-segregation problem. It was to gain some insight into this part of the problem that this work was undertaken. I) PROCEDURE At the outset of the work identification of the phases present and the elements concentrated in the freckle areas was undertaken. The electron micro-probe was used to determine general enrichment as well as -specific phase chemistry. Microradiography using cobalt and chromium radiation was also used on the freckle areas again to determine the elements involved in the segregation. X-ray diffraction patterns were made from portions of these same foil specimens. Optical microscopy and Knoop hardness measurements were also used

Jan 1, 1965

By Herbert Haas

CHEAP power is essential to large-scale mining and metallurgical operations, particularly where fine grinding of large tonnages has to be resorted to, as is the case with an increasing number of. mines treating disseminated copper ores, or low-grade gold ores, in the grinding of which considerable power is consumed. A comparison of the cost of generating power with Diesel engines and steam turbines and a brief discussion of the different factors governing the profitable use of either type of prime mover may be timely in view of the importance of. a cheap supply of power to the mining industry, and the prominence that has lately been given to Diesel engines in America. As many of our most important mines and largest producers of metals are located in sections where neither cheap coal nor water power is available, and the price of oil fuel is increased by long hauls, the Diesel. engine will. become a prime mover of increasing importance. This applies particularly to the Southwestern States, where an ample supply of cold water for condensing purposes is unusual and local climatological conditions interfere in securing the high vacua essential to high turbine efficiency. About 12 times as much water will be required for condensing purposes with steam turbines under general conditions as is needed for jacket-cooling of two-stroke, cycle Diesel engine; four-stroke cycle Diesel engines require only one-twentieth of the cooling water. Another condition favoring the use of Diesel engines is the high load factor at which most mine power plants operate. The influence of load factor on the cost of generating power is shown in Tables 1 and 2 and in Fig. 1.

Jan 7, 1916