Search Documents

By Dwane K. Johnson

INTRODUCTION The small mining company faces the dilemma of how to finance the development of its properties because it doesn't have the financial resources to pay for the development costs from its own funds and doesn't have the financial strength which will enable it to borrow the money needed for development. However, with a little imagination and the right set of circumstances, a financing can be arranged for the small mining company. This paper is an attempt to describe the fundamental steps on how such financing can be obtained --in other words--what is required to provide ad- equate capital at a low cost for developing and bringing a mining property into production. Innovative financing follows from this. The basic keys for building and maintaining a successful mining company are "the four M's": Management, Market, Mine, and Money. The writer believes their order of importance is as presented and each is a building block required in order to obtain the type of financing that best suits the borrower and lender. MANAGEMENT Management must be experienced, highly organized, realistic, able to take advantage of opportunities and fulfill promises. Mine financing requires a variety of skills and concentration on the part of management. Management should have a thorough understanding derived from experience in geology, construction, mining and metallurgy and environment, marketing, and financial matters. Management must realistically visualize what will be produced at what cost and truly evaluate the risks associated so the proper security can be correctly designed. Since risk cannot be eliminated, the objective of management is to identify the risk which will be present in a given venture and assess the level of that risk which will be acceptable to the firm. Generally the risk which is present is subject to little or no modification. After management identifies and measures the acceptable risk the firm will proceed in a way that it will be least exposed. Management must determine its long-term objectives and strategies within the context of a constantly changing world. This question must be addressed before examining the sorts of risk which affect the development and operation of a specific property. The compilation and interpretation of "hard data" by competent people is good to a point but the experience and instinct of seasoned individuals are the important factors in management choosing a long-term direction for the corporation. The attitude towards risk greatly affects the goals of the corporation. Major risks in a mining project can be placed in four categories: 1. Market - Price, demand, substitution. 2. Costs - Capital, operating, financial. 3. Regulation. 4. Taxation. The feasibility study for a project is the fundamental tool used in the management of risk. Management should employ an experienced team with an established and well-understood set of ground rules for the preparation and assembly of a feasibility study. The necessary degree of realism is built into all levels of the study and if the project is technically feasible, the company's hurdle rate is then used to discount the base-case project cash flow in order to determine its financial viability. Understanding the geological significance of the mineral deposit is critical. There should be strong interplay among the pure regional geologist, the detailed mine geologist, the mining engineer, and the metallurgical engineer. The collaboration among these different disciplines will yield a feasibility study that is more reliable. The essence of this work must be appreciated and understood by the executives planning the financing. It is vital that they thoroughly understand the risks and make the proper risk assessments to cope with our rapidly changing market environment. Projecting future revenue values is paramount and is subject to much estimation. One way to mitigate fluctuations and the risk of falling prices is by selling production for future delivery. MARKET Since mining is a worldwide industry it is mandatory that bankers be aware of what is occurring in the marketplace, both foreign and domestic. The marketing of the production is a very important facet when considering financing for the small

Jan 1, 1987

By P. E. Busby, C. Wells

FURTHER study of data used in determinations of 1—rates of diffusion of boron in austenite and 2—solubilities of boron in the a and phases of iron and steel' has provided an equation for the diffusion of boron in a iron. In brief, the previously published data were accumulated from the results of deboron-izing (and decarburizing) experiments carried out in the range of 700° to 1300°C. Diffusion coefficients (D?) for boron in austenite were calculated using the Grube solution of Fick's law. However, only solubility values were estimated from the discontinuous concentration-penetration curves, Fig. 1, which are characteristic of diffusion in two phases. Dr. Carl Wagner' has recently provided a solution for calculating D values from penetration curves of this type as follows: Cl1.1 —C8/C8- C11.1 = vp ? e ? erf (?) and D = ?/4?2 t where C11,12, C8, C11, and E have the values shown in Fig. 1. D is the diffusion coefficient, sq cm per sec; t is time of deboronizing anneal, see; and ? is a di-mensionless parameter. For a given diffusion experiment, the value of ? can be readily obtained by graphical solution from a plot of y vs the right-hand part of Eq. 1. D may then be evaluated from Eq. 2. The avplication of this solution to previously re-ported results, of which excerpts are given in Table I, permits the calculation of diffusion coefficients for boron in a iron. On the basis of these meager data, it is tentatively concluded that the diffusion of boron in a may be represented by the equation Da = 1011 e-92,900/RT where R is the gas constant, cal X 0C-1 X mol-1; and T is the absolute temperature, OK. Although the frequency factor, 106 sq cm per sec, is admittedly several orders of magnitude higher than expected, the value of Q, 62,000 cal per mol, appears reasonable and is, in fact, very similar to that for the self-diffusion of iron. It is pertinent to mention at this point that the value of Q obtained for the diffusion of boron in austenite by means of the Wagner solution is 19,000 cal per mol and is in excellent agreement with the value previously reported1 in the equation DT = 2xl0-3 e-21,000/RT [4] which was determined by the application of the Grube solution to other data. The fact that determined constants A and Q in the equation D? = A e-Q/RT were practically the same, independent of whether the Wagner or Grube solutions were used in determinations of D values, strengthened the authors' belief that the computed values of D (Table I) using the Wagner solution are reliable. The relative values in Eqs. 3 and 4 for the diffusion of boron in the a and ? phases, respectively, suggest that boron forms a substitutional solid solution in a iron and an interstitial solid solution in ? iron. The same tentative conclusion has been reached by McBride et al.3 on the basis of relative solubilities, atom diameter, and the size of the interstitial hole in a and ? iron. In connection with the data for test 14 listed in Table I, it is of interest to calculate the solubility of boron in a iron using the D value given by Eq. 3. As might be anticipated from the small movement of the interface in test 14, proper substitutions in Eqs. 2 and 1 give a low value, approximately 0.0004 pct B, at 850°C. Apparently at 835°C (test 13) it is possible to obtain 0.0018 pct B, and at 850°C only 0.0004 pct B into solution in the a phase before a second phase appears. These observations are consistent with the Fe-B constitution diagram proposed by McBride, Spretnak, and Speiser.3 References P. E. Busby, M. E. Warga, and C. Wells: Diffusion and Solubility of Boron in Iron and Steel. Trans. AIME (1953) 197, p. 1463; Journal of Metals (November 1953). 'W. Jost: Diffusion in Solids, Liquids, Gases. (1952) pp. 69-71. New York. Academic Press Inc. "C. C. McBride, J. W. Spretnak, and R. Speiser: A Study of the Fe-Fe2B System. Trans. ASM (1954) 46, p. 499.

Jan 1, 1955

By A. D. K. Laird, C. A. Johnson

Drag coefficients for a cylinder being towed through water at constant velocity and also at constant linear acceleration were measured. Drag coefficients for constant velocity towing showed reasonable agreement with previous results. The drag coefficients found for constant acceleration showed little correlation with either the magnitude of the acceleration or the value of the modulus AD/V2 and tended to be slightly lower than those for constant velocity. INTRODUCTION For the design of offshore structures, wave forces acting on immersed cylinders are of importance. Several investigations have been made of these forces and of how they can be predicted from wave theories. A paper by Crooke' summarizes much of the work already done. More recent research by the Wave Research Laboratory at the University of California, Berkeley, has added much to the body of data collected under ocean wave conditions, but has not found a reliable method of force prediction from basic theories of wave motion. The present investigation is part of a long-range program to collect data with which to test current and future theories and correlations. The immediate objective was to measure drag forces on a circular cylinder subjected to linear accelerations over the range of Reynolds numbers between l04 and 105 where the drag coefficient was a constant for constant velocities. In particular, the effect of constant acceleration from rest and from uniform motion was to be investigated. The effect of acceleration may be evaluated in terms of drag coefficients. A useful resistance coefficient, C, is defined by the equation for total drag force, D, exerted by the fluid on the cylinder. DT =CSp V2/2.........(1) The coefficient C may be plotted as a function of the parameter AD/V2. Here D is the diameter and S is the projected area of the cylinder. The acceleration and velocity are A and V. The mass density of the fluid is p. An alternate drag coefficient, C, corresponds to the total force of the fluid opposing the motion minus the virtual mass force. This drag coefficient, arising from the velocity of the fluid, may be compared with that for the same cylinder in uniform motion. The similarities and differences between this coefficient as calculated for accelerated and for steady motion are of interest in the inverse process of predicting drag forces on cylinders under arbitrary conditions. One serious difficulty in this method is the specification of the virtual mass which depends also on the flow picture about the cylinder.1 EXPERIMENTAL EQUIPMENT The experiment was performed in the combination wave and towing tank at the Engineering Field Station of the University of California. The tank is 200 ft long, 8 ft wide, and 6 ft deep. The tow carriage is externally propelled by a cable, and the velocity of the carriage is maintained essentially constant by means of an am-plidyne. For the purpose of the present investigation, the speed control of the amplidyne was modified by inserting a linear speed control rheostat, which was rotated at a constant speed by a small d-c motor equipped with a slip-clutch. Accelerations which were constant over a reasonable length of time were thus produced. The magnitude of the acceleration depended upon the speed of the d-c motor and the limiting speed of the carriage. Data were obtained for a range of accelerations of 2 to 10 ft/sec,2 and a range of velocities of 2 to 15 ft/sec. An accelerometer (Statham Type C-2-350) with a range of ±2g was attached to the carriage and connected to a Brush analyzer and penmotor. A continuous coil of wire with 10 turns/ft wound around a fixed wooden form extending for a distance of 40 ft along the length of the tank was connected by an electrical contact mounted on the carriage to a Brush analyzer and penmotor, to give the time-position history for each run. The test cylinder was machined from solid aluminum alloy, 17-ST. to a diameter of 1.250 in., and could be

Jan 1, 1957

By J. Tuzson, H. McL. Clark

Introduction Wear of pumps, wet cyclones and pipelines by suspended solid particles will impact their operating economies by limiting their useful life, reducing their reliability and increasing the amount of servicing. Erosion is considered a materials problem. The empirical solution consists of testing different materials in field applications. Such tests are time consuming and expensive. The prediction of localized wear rates, life estimates and appropriate material selection demand a better understanding of the slurry wear process in environments corresponding to those found in working equipment. Key words: Wear, Slurry abrasion, Coriolis erosion tester The slurry wear process Solid particles suspended in liquid generally move with the flow, but such particles deviate under the centrifugal and Coriolis forces encountered in pumps. The concentration of particles will increase, and a solid layer or bed may accumulate if the acceleration is directed against the flow passage wall. Large particles in dilute suspensions may impact the wall. dissipating their kinetic energies and removing material. However, particles smaller than about 11)(1 µm encounter great drag when moving through the liquid. and such par¬ticles have little kinetic energy left when they reach the wall. They cannot penetrate the stagnant boundary layers present in pumps. Therefore, small particles tend to form beds that slide along the wall, removing material by a process that could he called sliding erosion or wet abrasion. Wear by impact erosion from large particles in dilute slurries can he predicted by appropriate analytical and empirical methods. However, laboratory tests and calculations of wear rates that correspond to sliding erosion are more likely to correlate with field experience from wear under similar conditions. Specific energy Data from laboratory slurry erosion experiments have shown that the amount of material removed is approximately in proportion to the energy expended in the erosion process. Thus. the specific energy-energy expended in removing by erosion a unit volume of material-offers a simple, practical measure for the erosion resistance of materials. The concept was found to also apply to sliding erosion (Clark et al.. 1997). Generally, the study of the material-removal process in slurry erosion continues to remain a subject of research. Conceptually, if the flow velocities and accelerations could be calculated in a pump or pipeline, the energy dissipated at the passage walls could also be estimated, and the material removal rate would become known from the specific energy determined by laboratory tests. Slurry erosion testing Laboratory devices made specifically for slurry erosion testing may be grouped into the following: jet erosion devices, slurry pots and the Coriolis test machine (Pagalthivarthi and Helmly, 1992). Jet erosion devices use a slurry pump and nozzle to produce a high-velocity slurry jet that is directed at a flat material specimen, either perpendicularly or at some acute angle. Combined impact and sliding erosion are produced. but the processes cannot be separated, and the wear is difficult to assess except by weighing. Slurry pots consist of a cylindrical vessel containing the slurry of interest. In the slurry pot, test specimens are attached to a rotating arm. The samples are rotated in the slurry and experience local flow conditions that result in combined impact and sliding erosion. Wear depth measurements from detailed laboratory experiments have been successfully correlated with results from an analytical model of the process, and specific energy values have been calculated (Wong and Clark, 1995). In industrial use, the samples are weighed to determine the total material loss. Materials are ranked with respect to some reference material. The Coriolis erosion tester (Tuzson, 1984) consists of a rotor with a central cavity, from which two opposing radial passages lead out (Fig. 1). The slurry is fed to the centerof the rotor in one pass from a stirred overhead tank. The slurry particles are subjected to the Coriolis acceleration as they flow out through the radial passages. The material specimens are mounted into the radial passage walls facing the direction of rotation and are eroded by the particles that are pressed against them as they flow out radially. A narrow groove is worn into the sample within a few minutes, using a few gallons of slurry. Material removal is estimated accurately from profilometer traverses across the wear scar at different locations. The energy dissipated by the particles can be estimated, which leads to a well-defined experimental value for the specific energy.

Jan 1, 2000

By John Tabor, R. B. Spivey

INTRODUCTION Shafts have been sunk in a number of ways. By and large, however, most have been sunk by drill, blast, muck, and slip form methods. Most methods have provided satisfactory means of completion. The major drawbacks to the conventional methods have been: (1) un- reliable cost estimates for completion; (2) unreliable rates of progress; and (3) unreliable integrity of the shaft lining. Recently , engineering design studies have been done that strongly indicate, that, by the use of existing proven technology, shafts can be completed faster, more cheaply, safer, and with a better lining that has been possible using conventional methods. THE METHOD The alternate method utilizes the techniques of big hole drilling, shield excavator tunnelling, and pre-cast concrete lining. Big Hole Drilling - Drilled shafts have been completed for several years and I don't intend to address the technology, only to point out the use of drilled shafts is an alternate method of shaft sinking. Ideally, in a multi-shaft mine, the first shaft would be blind bored and cased to a size, whereby it can handle the muck from the excavation of subsequent shafts. It should be an 8' to 10' completed diameter. The cost of such a shaft in sedimentary formations compares favorably with a conventional shaft. However, it is much faster. For additional shafts, the role of the drill is to complete small holes (3' to 4') to the desired horizon. This is merely a hole to allow slashing to a larger size with easier muck removal from an underground drift connected to the previously drilled shaft. The smaller holes, just as the larger one, should be cased and grouted into place so as to assure a smooth open hole. SHIELD EXCAVATOR Theory and History - Tunnelling Shields have been in common use for several decades, and steam powered shields were used in England over 100 years ago. The purpose of the shield is to provide temporary ground support, protect personnel during operation, and to house the excavating equipment. The excavating equipment approach has varied from attempts at a continuous or semi-continuous boring machine, the tunnel mole, to a mechanical or hydraulic digging arm. The shield may or may not be an active digging element. If not active, the boring tool or cutter is advanced ahead of the shield to full diameter and then the shield is moved up to the cutter, or the shield advances along with the cutter. If active, the cutter opens a conical free face, and the shield is either jacked forward so that the leading edge spalls rock to the free face or poling plates are thrust forward to spall to the free face. In either approach, various methods of muck removal are incorporated into the shield. These may be gathering arms, conveyors, or with moles, hydraulic transport, where the cuttings are sufficiently fine. The efficiency of the muck removal system is critical as excavation is rapid and the system could easily become muck-bound. Application to Shaft Sinking- Technically, the use of a shield to sink vertical shafts is no different than a tunnelling operation. It is, in fact, less complex in most respects. Muck and water removal are difficult in a blind-driven shaft. However, where an underground opening is available for handling muck and water, rapid shield sinking is possible, using a pilot hole for muck and water to fall through and be removed from below. The Tabor Mining Shield is basically a caisson approach, whereby, the ground is only seen at the face or bottom. It does have hydraulic jacks to force it downward and has segmented cutting edges which allows forepoling ahead of the shield. The hydraulic excavator allows mining in the center of the face around the previously drilled hole with the forepoling segments slabbing to this free face. First, the drilled hole is cased and grouted. The machine is a tube with an excavator. The work cycle is as follows: (1) The casing and grout are cut at predetermined intervals by shaped charges. (2) The casing is crushed by the excavator, removed from the hole and hoisted to the surface. (3) The excavator then rips out a face around the hole by digging the formation in tension and breaking the pieces

Jan 1, 1982

By Jonathan Ludlow, Robert A. Jankowksi

Introduction A longwall operator can make few changes to increase output, significantly reduce respirable dust, and decrease power consumption. Reducing drum speed, and thereby cutting with increased pick penetration, is one. This article defines the benefits of deep cutting in terms of reduced dust production and power consumption. It also identifies the practical aspects of high pick penetration in terms of shearer performance and coal loading. Before examining some practical aspects of reducing drum speed and looking at the theoretical background, it is worthwhile to summarize what is meant by high penetration and deep cutting, and what potential benefits and pitfalls may be expected. Deep cutting (in the sense of high penetration rather than wide web) can be defined in one or more of the following ways: • Cutting with an average pick penetration distance higher than that used in the past. • Cutting with a pick penetration higher than the longwall operator would have used if the advantages of deep and slow cutting were not considered. • Cutting with a well-designed shearer drum below 40 rpm. All these definitions are slightly arbitrary. They are given to provide a basis for discussion and to make the point that any move towards deeper, more efficient cutting can result in operational benefits. The benefits of deep cutting appear in many different areas. The most noticeable benefit, provided suitable instruments are available, is the reduction of airborne respirable dust. During an experiment on a longwall in the Pittsburgh seam, a nearly four to one reduction in dust levels was seen when drum speed was halved. Not all studies have shown such a big reduction, but it seems that some benefit is almost always obtained when drum speed is reduced. Production rate and specific power consumption are also affected (in a positive sense) by reducing drum speed or increasing pick penetration. Although these changes may not be as spectacular as those in dust level, they contribute to the economic return of the longwall operation. Similarly, improved washability through fines reduction may have a beneficial economic effect. Cutting with shearer drums operating at lower speeds does have some possible deleterious impacts that an operator should be aware of. For example, cutting reactions - loads imposed on the picks by the coal being cut - will be increased as a deeper cut is used. Steps must, therefore, be taken to ensure the stability of the shearer and provide an adequate haulage effort. These increased cutting reactions also result in higher loads on the power transmission system (gearboxes, ranging arms, pick boxes, etc.) from the shearer motor(s) to the pick tip. These higher loads must be anticipated and provided for with the necessary hardware. In particular, extra haulage power must be provided with low drum speeds, since haulage effort required increases roughly in proportion with pick penetration. Because the drum will be rotating more slowly or will have fewer picks, the load on shearer components will also be more variable. If suitable, robust equipment is not used, this increased vibration will decrease reliability. Benefits of Deep Cutting Lower dust levels, decreased specific power consumption, and improved product washability are the most noticeable benefits of reduced drum speeds. Although the benefits will vary greatly with mining conditions and the type of coal, some examples of what can be expected are described below. Reduced Dust Levels Figure 1 shows principal results of a study on the effects of reduced drum speed conducted on a longwall in the Pittsburgh seam (Ludlow, 1981). This figure shows that average dust production was reduced by about 70% when drum speed was halved. By making some assumptions about such quantities as coal density, it is possible to apply this proportional reduction to the quantity of respirable dust liberated per ton of coal mined. When this is done, two kinds of results are obtained: • At 70 rpm, about 1 g (15 gr) of airborne respirable dust is created for every ton mined (roughly one part per million). At 35 rpm, only 0.28-0.37 g/t (3.9-5.1 gr per st) of coal mined become airborne respirable dust. • At 35 rpm, nearly four times the amount of coal may be mined before the compliance level is exceeded, compared with 70 rpm.

Jan 3, 1984

By G. T. Newell, J. G. Stone, V. M. Mejia

Introduction Advances in drilling have reached a point where large diameter cores can be recovered from "tight," or weakly indurated placer gravels. In such ground, core drilling can provide more reliable data regarding tenor than can be obtained using churn drilling or similar classical techniques. It can also provide metallurgical and geological information that is not available from samples obtained through alternate methods. In 1985, Coastal Mining Co, a subsidiary of M. A. Hanna, and Western Gold Reserves began to review a Tertiary placer deposit owned by San Juan Gold at North Columbia, CA, about 14 km (9 miles) northeast of Grass Valley. The deposit is one of the largest remaining unmined portions of the formerly extensive early Tertiary ancestral Yuba river system. It has been known since the 1850s, has been the subject of much technical literature, and has been the object of at least four previous drilling programs. The eastern one-third of the 6 km (3.7 mile) stretch of the channel between North Columbia and Badger Hill was partially stripped by large scale hydraulic mining in the late 1870s and early 1880s. Mining ceased in 1884 when the Sawyer Decision prohibited further discharge of hydraulic tailings into the Sacramento and San Joaquin Rivers. By that time, about 30 to 45 m (100 to 150 ft) of relatively low grade upper gravels had been removed over some 81 hm2 (200 acres). About 90 to 105 m (300 to 350 ft) of higher grade middle and lower gravels were left at least partially stripped. In 1914, a few churn holes were drilled along a widely-spaced line. In 1938-1939, Selection Trust conducted an extensive drilling campaign to evaluate the deposit. Particular attention was directed toward the partially stripped eastern portion. In 1968, the US Geological Survey drilled three churn holes in the eastern part of the deposit. The US Bureau of Mines conducted experimental mining and drilling in the Badger Hill area. In the late 1970s, Placer Service Corp. acquired a lease on the deposit. Between 1979 and 1984, Placer Service drilled 28 large diameter BADE (a German-manufactured machine) drill holes on the eastern portion of the deposit. The surviving records from the widely-spaced 1914 drilling program are fragmentary and the reported grade not well substantiated. The 1968 holes were drilled for scientific purposes. Again, drilling details are not available. However, detailed records for both the churn drilling program and the BADE program were available and formed the basis for the initial evaluation of the property. Geology The geology of the auriferous Tertiary gravels of California have been described by Whitney (1880), Lingren (1911), and, more recently, Yeend (1974). In general, the Tertiary gravels in the North Columbia area occupy a broad channel cut into pre-Tertiary igneous and metamorphic rocks. The upper, or white gravel is overlain conformably by volcanic tuffs and volcaniclastic rocks. A middle gravel is characterized by the presence of silicified and carbonized wood. A lower blue gravel unit has relatively coarser cobbles and contains a higher proportion of igneous and metamorphic cobbles than the other units. The upper gravel consists of interbedded pebbly sand and silty, or clayey sands with prominent cross bedding. Most of the pebbles are well rounded and consist mostly of white vein quartz and quartzite. The upper unit is moderately well compacted. Exposures in the walls of the old hydraulic mine pits stand at 45° and 50° angles. The gold content of the unit is well below an economic cutoff. The middle gravel - included with the upper unit by Yeend (1974) - is coarser grained, with carbonized wood, and 75 to 100 mm (3 to 4 in.) cobbles of metased-imentary and metavolcanic rocks in a sandy matrix containing abundant lithic fragments. The upper contact appears to be conformable, but the lower portion of the unit appears in places to consist of reworked lower gravels. The unit contains less clay than the upper unit and is somewhat more friable than the underlying lower gravels. The gold content, while somewhat higher than the upper level, is too low to be of ore grade. The lower gravel averages between 30 to 45 m (100 to 150 ft)

Jan 9, 1988

By Marc C. Bétournay

A mine production locomotive, the world?s first applied industrial fuelcell vehicle, has been designed and tested in the context of a number of proof-of-concept projects supported by the Fuelcell Propulsion Institute, an international consortium of stakeholders dedicated to facilitate the commercialization of fuelcell vehicles. The locomotive?s simple electric-powered motor and controller system was ideal to study a number of requirements related to fuelcell power application: hydrogen power plant operation, underground hydrogen use, risk and regulatory requirements, production performance versus the conventional battery-powered locomotive version under measured duty cycles and production requirements, power plant design improvements for other fuelcell vehicles, and cost-benefit aspects. This article will focus on the design and testing aspects of this project which has led to successful performance demonstrations of the vehicle under surface and underground conditions at two mines. It will also present information on the creation of a safe, reliable and efficient power plant which fulfilled its design objectives, including longer productivity time for a given vehicle fill-up.

May 1, 2003

By B. M. Loring, I. R. Krame, F. M. Walters

NO observations on the mechanical properties of man-manganese alloys have been published since pure manganese became readily available, either distilled manganese or electrolytic manganese. The purpose of this investigation was to determine whether iron-manganese alloys are essentially brittle, as reported by Hadfield,' or whether the lack of ductility observed was due to impurities in the alloys that he studied. Experimental Methods The alloys were melted in a roo-lb. high-frequency induction furnace with a magnesia crucible, using ingot iron and electrolytic manganese. The heats were split, the first giving six 18-lb, ingots and the second three. Each successive heat was higher in manganese and about 0.3 per cent silicon was added as ferrosilicon. The alloys forged without difficulty and were given a 9' per Published by permission of the Navy Department. Manuscript received at the office of the Institute June 28. 1941. Issued as T.P. 1369 in Metals Technology. September 1941. * Division of Physical Metallfirgy Naval Roscarch Laboratory. Anacostia Station Washington. D. C. 1 References are at the end of the paper. 4 cent reduction to ¾-in. rounds. The chemical analysis of the alloys is given in Table I. The forged rods were homogenized for 224 hr. at r150°C. (2100°F.) and air-cooled. The alloys were tested in two conditions of heat-treatment: (I) as normalized from 870°C. (1600°F.) and (2) normalized from 870°C. and tempered for I hr. at 540°C. (Iooo°F.) and air-cooled. The tensils specimens were 0.505 in. in diameter with a 2-in. gauge length, and the impact specimens were v-notch Charpy. The proportional limit and yield strength were taken from the stress-strain diagrams drawn by a Peters recorder. The data obtained from the tensile test and the hardness measurements are given in Table 2. Specimens that were normalized are designated by N; those that were normalized and tempered by T. Alloy G was tempered after forging (D).

Jan 1, 1942

By M. J. McPherson

Having worked on the thermodynamics of air/liquid-water mixtures passing through the surface fans of deep mines, I find this paper of great interest and congratulate the authors on producing it. There are two matters, however, deserving discussion. First, the authors have described the classical theory of fog formation with fogging occurring at "supersaturation." In fact, the process of fog formation begins well below 100% humidity. The more strongly hygroscopic nuclei in the atmosphere will attract water molecules and begin to grow at a relative humidity perhaps as low as 70%. The process is a dynamic one with both condensation and evaporation taking place throughout the mixture on microscopic liquid surfaces. As saturation is approached, the rate of condensation accelerates rapidly producing the familiar reduction in visibility. Hygroscopic nuclei are present in all natural atmospheres and under the appropriate conditions of pressure and temperature, will produce clean fogs. However, if the air is polluted by particulates from combustion or other processes then the resultant coagulation with growing liquid particles may produce dense (and sometimes photochemical) smogs. It is this process that is likely to occur in underground mines when moist air is cooled below dew point. Second, the authors have summarized very well the individual measures that might alleviate the problem. In particular, I agree that neither heating nor refrigerating the air, by themselves, provides a satisfactory solution. However, there is a combination of these that provides a neat and effective control of fog formation. This involves a small, self-contained refrigeration unit within a duct but without the usual external heat rejection facility. The air is cooled below dew point and, hence, dehumidified on passing over the cold evaporator coils. The heat from the condenser is rejected back into the air down-stream from the water eliminator, as shown in [Fig. 1]. The duct configuration can be designed to create good mixing at the outlet.

Jan 1, 1984

By B. M. Loring, I. R. Krame, F. M. Walters

NO observations on the mechanical properties of man-manganese alloys have been published since pure manganese became readily available, either distilled manganese or electrolytic manganese. The purpose of this investigation was to determine whether iron-manganese alloys are essentially brittle, as reported by Hadfield,' or whether the lack of ductility observed was due to impurities in the alloys that he studied. Experimental Methods The alloys were melted in a roo-lb. high-frequency induction furnace with a magnesia crucible, using ingot iron and electrolytic manganese. The heats were split, the first giving six 18-lb, ingots and the second three. Each successive heat was higher in manganese and about 0.3 per cent silicon was added as ferrosilicon. The alloys forged without difficulty and were given a 9' per Published by permission of the Navy Department. Manuscript received at the office of the Institute June 28. 1941. Issued as T.P. 1369 in Metals Technology. September 1941. * Division of Physical Metallfirgy Naval Roscarch Laboratory. Anacostia Station Washington. D. C. 1 References are at the end of the paper. 4 cent reduction to ¾-in. rounds. The chemical analysis of the alloys is given in Table I. The forged rods were homogenized for 224 hr. at r150°C. (2100°F.) and air-cooled. The alloys were tested in two conditions of heat-treatment: (I) as normalized from 870°C. (1600°F.) and (2) normalized from 870°C. and tempered for I hr. at 540°C. (Iooo°F.) and air-cooled. The tensils specimens were 0.505 in. in diameter with a 2-in. gauge length, and the impact specimens were v-notch Charpy. The proportional limit and yield strength were taken from the stress-strain diagrams drawn by a Peters recorder. The data obtained from the tensile test and the hardness measurements are given in Table 2. Specimens that were normalized are designated by N; those that were normalized and tempered by T. Alloy G was tempered after forging (D).

Jan 1, 1942

IN the ordinary course of treatment of pyritic ores or concentrates, whether for the recovery of copper, tin, gold, or any other contents, the removal of some or all of the sulphur contents is almost invariably the first step. The practice is to roast the ore in a current of air so that the sulphur is oxidized to sulphur dioxide and the metal constituents are changed from sulphides to oxides. If it is feasible to produce sulphuric acid in conjunction with the treatment process, the sulphur can be usefully employed; but if this is not possible, then the sulphur contents not only become a dead loss, but tend to be a nuisance as well in the surrounding neighbourhood.The problem of avoiding this disadvantage and making economic use of the sulphur has had attention from the early days of scientific metallurgy, and in the past few years very special efforts have been made both to drive part of the sulphur as such from pyrites and recover it, and to replace the sulphur in the S02 by carbon, with the production of carbon dioxide and precipitation of the sulphur. So far economic success does not appear to have crowned these efforts. In the case of pyrites, of the two atoms of sulphur which go to make the FeS2, the first is very lightly held. Dissociation is effected by heat alone at a temperature which is given by Peters as 5160 C.Now, sulphur and hydrogen will unite or burn together, with the production of sulphuretted hydrogen. It occurred to the writer that, if hydrogen was passed over pyrites heated to the above temperature, this first atom would be snatched away by the hydrogen, and could possibly be economically recovered.The first step consisted of some simple laboratory experiments. From these it was found that this first atom was readily removed, and that the extraction was practically complete. Further, the reaction took place at a low temperature, a dull red heat being sufficient.

Jan 1, 1923

By H. Scott

IN a paper read before the Institute of Metals, Rosenhain 1 described a new type of furnace designed primarily for the thermal analysis of metals by the inverse-rate method and used by him in the metallurgical department of the National Physical Laboratory with considerable success. In his discussion of this furnace, Rosenhain pointed out some difficulties met with in its operation, such as uniformity of rate of heating or cooling being inadequate for the degree of accuracy aimed at. To overcome this difficulty, he suggested, in place of motor propulsion, a gravity drive controlled by a"hydraulic cylinder with a relief valve whose width of opening can he regulated to allow of any desired rate of motion." The authors, in constructing a thermal-analysis furnace of Rosenhain's type, have therefore followed this suggestion and also added certain features that increase somewhat the convenience and simplicity of its operation. Requests for information regarding this furnace and the highly satisfactory results obtained from its use, it is believed, justify describing its construction and operation in sufficient detail to make possible its duplication or improvement. Description of Furnace.-The details of the furnace construction are shown in Fig. 1, which is drawn to scale. The heating tube is of /4 in. (6.35 mm.) wall "alundum" heated at the upper end by seventeen turns of 0.52 mm. platinum wire, which is coated with "alundum" cement supplied for this purpose. The cement coating is essential when a high temperature (over 1000° C.) is required, as it prevents hot spots with the resulting burning out of the heater. This platinum-wire winding, unlike "nichrome," is entirely satisfactory for temperatures of at least 1000° C. It has been maintained at that temperature continuously for 2 months and shows no signs of deterioration. This temperature is maintained by a current of 5 amp. drawn from 30 volts potential, so its necessarily continuous operation is quite economical.

Jan 8, 1919

By Fred W. Gordon

The North Chicago Rolling-Mill Co., of Chicago, have four furnaces at South Chicago, built during 1881. Each furnace is 20 feet diameter of bosh, and 75 feet total height, the hearth being 11 feet diameter, and the top 15 feet 4 inches. They are blown through seven 7-inch tuyeres placed 5 feet 6 inches above the hearth level. Cubical contents, 15,000 cubic feet. Each furnace is supplied with 2 blowing engines, 84-inch blowing cylinder, and 36-inch steam cylinder, and 54-inch stroke. These two have one jet-condenser worked by an auxiliary pump. The blast for each furnace is heated by the aid of three Whitwell fire-brick stoves, each 21 feet diameter and 60 feet high. In the batteries of boilers furnishing steam for four furnaces there are 36 double-length boilers, 76 feet long, 48 inches diameter and 16,000 square feet heating-surface. The plant was designed to produce 3600 tons per week, or 900 tons for each furnace, and was fitted as above with such engines, boilers, and hot-blast stoves as were thought requisite for that production ; hence, calculated from the present make—not the least remarkable feature is the economy of construction—there are less engines, less boilers, and less hot-blast power for the iron made than at ally plant within my knowledge, or it could to-day be duplicated for less money per ton of iron made per week than could any plant now in existence, whether using fire-brick or cast-iron stoves for heating the blast. This is a practical demonstration of my assertion in a paper on fire-brick stoves, read before the Institute," that the application of fire-brick stoves in construction really lessens the first cost. In view of the statements just made, I must be most emphatic in assuring the members that nothing is strained or overburdened.

Jan 1, 1886

By Clemens Jones

The peculiar property of the lode-stone has been known for many centuries. It was very early observed that, by contact with it, iron is magnetized; and it is needless to dwell here upon the history allied to this discovery, now constituting the wonderful science of magnetism. Deposits of magnetic ore are distributed all over the world. The ore is sometimes very rich and comparatively free from gangue, sometimes sparsely disseminated in the native rock. These bodies of lean ore often cover large areas. Modern ingenuity has utilized the principle of the magrtet, originally derived from the ore itself, and really belonging to the time of Plato and Theophrastus, to separate the magnetic ore-particles from the rock which has previously been crushed. This process recovers the ore in the state of a concentrate very rich in iron and, in some cases, desirable for Bessemer purposes. The magnetic separator is no longer a novelty. It was only at the last meeting of the Institute in New York, that Messrs. Birkinbine and Edison (Trans., xvii., 728) and Mr. Robert Cook (Ib., 599) described some of' the principal designs of apparatus in use. E. S. Dana classifies the natural magnetic minerals in the following order: magnetite, pyrrhotite, franklinite, almandite and minerals " containing considerable FeO." Of these, ferroso-ferric oxide, or magnetite, is the type. The proportionate amount of ferrous oxide essential to magnetize the compound, is not established. Ferrous oxide appears to be the only magnetic oxide of iron. Tissandier (Compt. Rend., lxxiv., 531) produced ferrous oxide in the anhydrous state, by the action of carbon dioxide on pure iron at a red heat, according to the equation : Fe + CO2 = FeO + CO. This is black, crystalline and magnetic. Heated to redness in the air, it is converted into Fe3O4. Sidot (Compt. Rend., lxvii., 175) has also produced an oxide exhibiting magnetic polarity, by heating ferric oxide to a temperature

Jan 1, 1891

By H. W. Camp

The past year has had no revolutionary change in the process of petroleum refining, although there have been improvements and developments in practically every phase of operation, due largely, perhaps, to economic pressure. A few years ago, a Mid-Continent skimming plant using 35 gravity crude obtained 28 per cent. gasoline and naphtha; 9 per cent. kerosene, and the remaining yield gave gas oil and fuel oil. Figuring manufacturing cost, but not including sales, transportation, overhead and depreciation, the realization was $1.45 per bbl. of crude. Today, a thoroughly modern and up-to-date skimming plant, having all latest equipment, shows a realization of $1.74 per bbl. on the same basis. Thus it is evident that progressive developments have taken place in the last 2 or 3 years when potential earnings have been increased about 20 per cent. The average refiner is compelled to exert his utmost ingenuity in order to operate with a profitable margin. He has two factors to consider—recovery of maximum yields of profitable oils and the reduction of operating expenses. For the past 8 years, the margin between price of gasoline and cost of crude has grown narrower, constantly. Both these prices are determined by external conditions and not by the refiner himself, but the refiner might have been able to increase that margin had he more carefully controlled his output, limiting his production of light oils to current consumption. Be that as it may, this same economic pressure has given impetus to progress in improvements of primary distillation equipment, including stills, towers and heat-saving devices. A pipe still and bubble cap fractionating equipment can now replace a battery of shell stills, giving better yields, better control, elimination of rerunning and lowering of operating costs. The most recent pipe stills are being equipped either with radiant-heat furnaces or air-preheating systems and improved heat exchange equipment. These can be operated on a fuel consumption as low as 1 per cent. of the charge as compared to a fuel consumption of

Jan 1, 1929

By John C. McDonald

The amount of published literature in the field of mechanical properties of magnesium alloys is not great; particularly with respect to rolled alloys. Haughton and Prytherchl have summarized most of the articles to 1937, including their own work on sheet. Although many papers can be found dealing with the technology of magnesium and the alloys now in commercial use, there does not seem to be any work dealing exclusively with the variation in properties of rolled alloys as caused by changes in alloy composition. Comparison of published data reveals considerable discrepancies in properties of alloys of the same composition, probably due to differences in the details of working and heat-treating followed by different workers. The purpose of this paper is to study the properties of alloys in conditions where the effects of such differences have been eliminated. It is well known that the strength and ductility of metals are functions of the working and heat-treating they have received, as well as of their chemical composition. In comparing alloys with respect to these properties, therefore, all possible combinations of the variables of the fabrication procedure ought to be carried out on each alloy. One or more of these treatments mill produce the highest ductility, or strength, or both. These highest properties would then be selected as characteristic of each alloy, and alloys would be compared on the basis of those properties. Thus a separation of the effect of composition and that of workiiig and heat-treating would be achieved. In this paper, however, a compromise with completeness has been made by carrying out a restricted number of treatments. As a consequence, the properties obtained are not necessarily the best that the alloy is capable of exhibiting; but since all alloy3 have been treated by the same method, the results ought to be comparable. The elements selected for the study comprise a class that will alloy easily with magnesium, and that are reblatively cheap. Each was added in different amounts, up to a limit that varied with the element.

Jan 1, 1940

By A. T. Peters

DURING an investigation concerning the use of large nozzles for pouring of low-carbon rimming steels, the practical details of which were reported elsewhere, it was noticed that increases of nozzle sizes were generally followed by improvements of slab surface quality. It is the purpose of this note to present the data available and some tentative opinions about this phenomenon. The problem of correlating pouring rates with the metallurgical performance of steels is complicated by several factors. The two most important are; 1) Variations of the steel pouring rate throughout a pour, these being due to the nozzle erosion, which enlarges the bore area, and the dropping ferrostatic pressure in the ladle. The pouring rate is lowest towards the end of a heat when the ladle is nearly empty. 2) Variations in the ingot pouring time due to the different ingot weights poured. Consequently, the use of the average pouring rate of a heat may not be exact, but in large-scale production it is not practical to treat ingots separately according to their pouring rates. The variations of ingot weights and sizes may be overcome by considering the rate of rise of steel in the mold. Hence, the average rate of filling of molds throughout the pour of a heat, inches per second, will be used in this note. Every ingot rolled at Inland is rated at the blooming mills prior to hot machine scarfing as being good, fair, or poor for cracks, scabs, and checking. The latter term refers to the network of breaks or lines associated with overheating or burning in the soaking pits. It is also known that thin-skinned steel is prone to show this defect if the subsurface blowholes are exposed during reheating or rolling. The experiences with the surface quality of slabs rolled from ingots poured at different rates are presented in Fig. 1. The data represent nearly 11,000 ingots. It appears that all three types of defects are in-

Jan 1, 1960

By John C. McDonald

The amount of published literature in the field of mechanical properties of magnesium alloys is not great; particularly with respect to rolled alloys. Haughton and Prytherchl have summarized most of the articles to 1937, including their own work on sheet. Although many papers can be found dealing with the technology of magnesium and the alloys now in commercial use, there does not seem to be any work dealing exclusively with the variation in properties of rolled alloys as caused by changes in alloy composition. Comparison of published data reveals considerable discrepancies in properties of alloys of the same composition, probably due to differences in the details of working and heat-treating followed by different workers. The purpose of this paper is to study the properties of alloys in conditions where the effects of such differences have been eliminated. It is well known that the strength and ductility of metals are functions of the working and heat-treating they have received, as well as of their chemical composition. In comparing alloys with respect to these properties, therefore, all possible combinations of the variables of the fabrication procedure ought to be carried out on each alloy. One or more of these treatments mill produce the highest ductility, or strength, or both. These highest properties would then be selected as characteristic of each alloy, and alloys would be compared on the basis of those properties. Thus a separation of the effect of composition and that of workiiig and heat-treating would be achieved. In this paper, however, a compromise with completeness has been made by carrying out a restricted number of treatments. As a consequence, the properties obtained are not necessarily the best that the alloy is capable of exhibiting; but since all alloy3 have been treated by the same method, the results ought to be comparable. The elements selected for the study comprise a class that will alloy easily with magnesium, and that are reblatively cheap. Each was added in different amounts, up to a limit that varied with the element.

Jan 1, 1940

By R. Moldenke

Unlike the furnaces employed in the reduction of ores to mattes and metals, the foundry-cupola has only melting to do. This looks simple enough; and its development has progressed through centuries by cut-and-try methods; but there is still much to be learned concerning the causes which underlie the production of defective castings, even from first-rate materials. There are other differences between the smelter and the cupola. The former is usually, the latter rarely, run continuously. In the smelter there is some chance for experiment, since rich slags can be treated again; while in the cupola the metal must be good for casting from the first tap to the drop of the bottom ; otherwise, heavy damage will result. In the smelter, the object sought is the production of metal or matte with a minimum loss in the slag; whereas, in the foundry, a variety of castings of different compositions must often be run from the same heat, one after the other. Experience with the cupola has shown that the melting should take place, throughout the heat, in one zone, the zone of maximum temperature, which is clearly marked on the cupola-lining by the severe cutting action of the slag and iron oxide. The design and operation of the cupola must conform to this condition. This conformity having been secured, other considerations resolve themselves into those details of construction which will involve least trouble and cost of repairs. The very high temperatures required to melt iron and steel in the cupola, leaving the metal sufficiently superheated for casting, render the use of water-jackets inadvisable. It is always possible to patch up the lining at the melting-zone between daily heats; and the rest of the lining is seldom affected enough by abrasion and chemical action to require replacement oftener than once in nine months. The cupola is practically a shell of steel, lined with the proper refractory material, and provided with a sct of tuyeres for the admission of air to burn the fuel and melt the metallic

Jan 1, 1913