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By Bob J. Rawicki

TYPES OF WET DUST COLLECTORS Basically, there are two types of wettype dust collectors. One is mechanical, incorporating pumps, motors, fans, sprays, filters, or flooded beds. These come in many forms, but their operating principle is basically the same. The screen or flooded bed is wetted by a series of sprays. Polluted air is drawn in through a fan; here the dust particles impinge on the screen and are flushed off into a settlement tank. The air then passes through spray eliminators to the atmosphere. A great deal of time and money has been spent developing this type of collector. In laboratory tests where clean water at constant pressure, air at constant volume, and dust at low but constant mass to volume were applied, these collectors gave high percentage dust collection efficiencies. Theoretically and technically, they are very good. In practice, however, under rugged mining conditions where there are no constants, and dust load, particle size, water pressure, and volume of air vary with time, their efficiency varies dramatically. Some of the most common problems are: 1. clogging of screens 2. blocking of nozzles 3. replacement. of damaged pump stators Any of the above result in loss of air flow and dust collection efficiency. Generally, there are so many mechanical moving parts, something is always breaking down. This type of collector is very expensive in parts and labor to maintain. The second type of dust collector, which I personally developed, with the help of the British National Coal Board of Great Britain, has none of the above mentioned problems and it works on a completely different principle. My collector, the Mark III Precipitaire Wet-Type Dust Collector, (an improved version of previously successful models) has no internal moving parts, no flooded beds, screens, spray nozzles, etc. It is highly efficient - 99.8% efficiency on total collected dust and 97% on respirable dust. The pressure drop through the collector is constant, hence, air flow remains constant. There is no maintenance expenditure, be it labor or parts. The only maintenance that there is, is desludging. This, of course, varies with the dust load. With a low dust load, desludqinq may only be required every 4 weeks, but this is a very simple operation that can be handled by unskilled men and it takes only a very short time. My collector is powered by a ventilation fan that is placed on the clean air side for a longer fan and motor life. Dust is extracted solely by water action. The collector's operating principle is: Dust laden air is drawn in along ducting into a tapered scrubber section, shaped to create a self-induced curtain of water. This action washes the dust particles from the air and the collected dust settles into the bottom of the tank. Cleaned air is then exhausted via a series of spray eliminators where it is completely dried and on into the atmosphere. These collectors have been sold in Great Britain, to the National Coal Board, and throughout Europe. They have been working very well for many years now. GENERAL APPLICATIONS Dust is very dangerous; silicosis and black lung kill and disable many workers each year. Many of the mine explosions can be attributed to coal dust. By the use and application of dust collectors, these deaths and disasters can be avoided. There really is no excuse. However good and efficient a dust collector may be, it's positioning and installation arrangement is critical. The collector may remove 99% of dust supplied to it, but it takes the skill and experience of a dust control engineer to design and fit ducting, hoods, etc. to remove polluted air at source and deliver it to the collector. In some cases this is very simple; in the case of longwall machines, it is very difficult. In general, dust collectors should be used to collect dust at every point of dust generation. I have noted a great reliance on dilution as a means of dust control. I personally think this is a dangerous and unreliable method. As mines become deeper and the use of larger and more productive equipment becomes more common, reliance on dilution will be impossible and dust collectors will be essential.

Jan 1, 1982

By C. Kremer Bain

UNDER present economic conditions the necessity of mechanizing the mines of our country has become a very important problem. More and more mines are looking toward increased or complete mechanization of their operations, so in the writer's opinion it is quite pertinent that some of the associated problems that accompany such a program be pointed out. The following generalization on attempting a comparatively slight change in mining methods at the Southeast Missouri properties of the St. Joseph Lead Co. is an attempt to point out to anyone faced with a mechanization program that much time and C. KREMER BAIN, Member AIME, is Assistant to General Manager, St. Joseph Lead Co., Bonne Terre, Mo. AIME Sun Francisco Meeting, February 1949. TP 2743 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Dec. 27, 1948; revision received Oct. 13, 1949. effort can be saved, and many headaches avoided, if company policy can be established whereby everyone in the organization from the top down is desirous of improvement and is working towards the common objective. Any changes in mining methods, machinery, and so on, must entah coordination of all the other phases of mining as they all have to mesh as accurately as a smooth running chain of gears if satisfactory results are to be obtained. The purpose of this paper is to point out some of the problems which :have to be surmounted when any mechanization program is attempted, rather than to go into the mechanical details and actual operating data of the equipment employed. After all, machines have to be adapted to best suit the local conditions, so no general formula can be given that would apply to other properties. Although St. Josepll Lead Co. has been mechani-cally-minded for a number of years, as shown by the fact that over 110 million tons have been loaded mechanically since 1922, any further mechanization or changes always present a problem. When our ore bodies were of great height and area the breaking and loading were developed to a high degree of efficiency according to known standards. But in recent years, when our ore bodies became smaller and more scattered, the amount of tonnage we were able to break was not sufficient to supply our shovels with enough rock to keep them working to capacity. Our first step in attempting to rectify this situation was to develop a technique of moving the shovels to the broken ore in various stopes, which sometimes entailed a move each shift. A shovel is not producing when being moved, so within the last two years we have been endeavoring to devise some means of getting more rock broken in these smaller stopes and it was then that we became interested in jumbo drilling. To illustrate the fact that mechanization is a process of a chain of developments rather than a stroke of genius can best be shown by a general summary of conditions and obstacles which were encountered in our present attempt to establish template drilling of a standardized round. Our first jumbo was more or less the conventional type with two machines mounted on a crossbar. This arrangement proved to be impractical to duplicate satisfactorily the regular routine of slabbing in a more or less radial pattern as carried out by column drilling. One machine was either too close to or too far from the face and there was another disadvantage in that one drillman was frequently delayed while waiting for his partner to bottom a hole. Our next step was to eliminate the crossbar so the machines could operate independently of each other, but we still encountered the same objections mentioned above except that one drillman was free to drill three holes instead of one in a vertical plane independently of his partner. Time studies while trying to adapt the jumbo to the regular slabbing practice as accomplished by column drilling, revealed that twice as much time was spent in maneuvering the jumbo into position and aligning the holes as was consumed in drilling.

Jan 1, 1951

By Jean-Jacques Prompsy, J. S. Rinehart

Some time ago a study was initiated at the Colorado School of Mines in an effort to arrive at a better understanding of the stress fields developed within drill bits under dynamic loading and the influence that these stress fields could have on the failure of the bits. Particular emphasis has been given to the influence that bit shape has on the establishment of highly localized, potentially destructive stress inhomogeneities within the bit. The study has been divided into three phases involving three velocity regimes: impacts at very low velocities, 0 to 20 ft/sec, a velocity range in which dynamic effects are just beginning to be found; impacts at velocities ranging from 20 to several hundred ft/sec, a range commonly encountered in practical drilling operations; and impulsive loading through the detonation of explosives, a region in which the dynamic effects are greatly exaggerated and made more identifiable. The ultimate objective is to provide basic data which will enable the mitigation through judicious design of the frequency and severity of these transient concentrations of intense stress, thereby prolonging and increasing drill efficiency. This paper presents results of the third phase of this study — the development of a better understanding of the dynamics and mechanics of failures caused by transient stress-wave interactions. The impulsive loading or transient waves in these experiments were generated by detonating small explosive charges placed on the ends of Plexiglas cylindrical rods terminating in truncated cones. Under such intense loading, dynamic effects are greatly exaggerated and, hence, made more identifiable. The explosion generates a high-intensity transient stress disturbance which moves through the rod, reflects from surfaces which lie in its path, and establishes momentarily narrow regions of high concentrations of stress where failure of fracture may occur. The peregrinations of these waves, their interactions and their effects have already been described in the literature (see, for example, Refs. 1 and 2). The locations and extent of the regions quite apparently are strongly dependent on specimen shape, making it possible through judicious shaping of specimens to relocate the highly stressed regions, minimize their extent, or eliminate them entirely. These experiments were designed, first, to identify more clearly the vulnerable regions in geometries relevant to drill-bit design and, then, to modify these regions in a predictable way by changing specimen shape. Through this latter process, it is anticipated that better bit design might evolve. The specimens used in the experiments were cylindrical rods, 1 1/2 and 1 1/8 in. in diameter, ended by a truncated cone, with the cone angles ranging from 45° to 130°. Upper cone face diameters were 1/2, 3/4, or 1 in. A small plastic-covered electric blasting cap detonated on the axis of the specimen produced the spherical shock wave. The detonators used were Olin Mathieson No. 6, which induced into the rod a transient saw-toothed stress disturbance of about 3-microseconds duration, corresponding to a wave length of 0.3 in. in the Plexiglas. Five distinct systems of fracture were observed: (1) along the axis of the rod, a lower system of fractures, the cracks focusing around one point for specimens having cone angles of 80° and 90° and, as the cone angle was increased, spreading downward; (2) just above these fractures, an upper system of fractures composed of horizontal cracks observable in some specimens but not of any appreciable extent; (3) a third system of axial cracks still above the latter, observable in the small cone-angle specimens; (4) in the upper part of the specimen, a system of radial fractures extending from the blasting point; and (5) circular spalling due to reflection of the wave on the boundary of the cylinder. Only the first three systems, the axial fractures, are examined here. LOWER TENSILE FRACTURE SYSTEM The first, lower system of tensile fractures, shown in the photograph of Fig. 1, was observed in almost all of the specimens. It was due to the reflected tensile wave coming in from the boundary of the cylinder as shown in Fig. 2. The reflected wave was cylindrically convergent and focused

By H. L. Riley, B. W. Gandrud

A froth flotation unit was added to the DeBardeleben Coal Corporation's Empire washer at Empire, Alabama, in the fall Of I94O following an investigation of possible ways and means of recovering coal values from settling-tank overflow or slurry. Officials of the DeBardeleben corporation hoped that some practical and reasonably economical way could be found for recovering fine coal from overflow water, although there had as yet been very little progress along this line in the Alabama coal fields. The Bureau of Mines was asked to cooperate in the project through its Southern Experiment Station, at University, Alabama. This paper presents a brief description of the process and the installation that resulted from the investigation of the overflow water or slurry problem at the Empire mine. The experimental data include' as a Part of this Paper were obtained during the course of the investigation in 1940 Since that time, however, the flotation unit has been running more or less continuously without significant operational changes and is still Performing pretty much in accordance with the data obtained in 1940. Flowsheet at Empire and Type of Coal Treated The Empire mine is developed on the Black Creek bed near the the border of the Warrior field. The coal in this bed is extremely low-ash and although there is an almost complete absence of bone coal and partings, the bed is so thin that, with mechanized mining, a high percentage of rock from roof and floor inevitably gets into the run-of-mine coal. The thickness of the Black Creek bed throughout the part of the Warrior field in which it is mined extensively is rarely over 30 in. and probably would average about 25 in, the coal is of the high-volatile A rank and of excellent quality, but as a result of admixture of roof and floor material, the run-of-mine product usually contains I ; to 20 pct rock by weight. he following is a proximate analysis on an "as-received " basis of a typical seam sample from the Empire mine:' moisture, 3.0 pet; volatile, 34.6 fixed carbon 58.3; ash, 4.1 This sample shelved 15,070 Btu per pound and 0.9 pct sulphur on a moisture-and-ash-free basis. In the Empire preparation plant the lump sizes above 3 in. are removed and hand-picked on a Marcus conveying Screen, the 3 to ? -in. size is removed by vibrating screens and washed in two Montgomery single-cell jigs, and the ?-in. to o slack is cleaned on four diagonal

Jan 1, 1949

By J. H. Craig

As a result of information that the Navy Department was questioning the advisability of raising the maximum allowable exposure temperature for zinc-coated steel parts from 500 to 750 F., it was decided to make a series of high temperature tension tests on several types of uncoated and zinc-coated steels. The objective was to determine at what temperature intergranular penetration of galvanized steel by zinc might take place under conditions of rapidly applied stress and temperature. Tests were conducted in the range from room temperature to 1400 F. In this range, low-carbon and high-carbon steels did not suffer any adverse effects due to zinc coatings. In fact, at 1400 F. the zinc coatings appeared to have a beneficial effect. In stainless steel, the coating had no effect up to 1000 F. Above this temperature, the zinc coating appeared to cause a marked loss in ductility and a less pronounced loss in strength. There was no visible evidence of zinc penetration of the stainless steel structure, but it was assumed that this was the cause of the adverse effect.

Jan 1, 1945

By Orville R. Lyons

THE relative performance of coal washing equipment, or the effectiveness with which any type or make of equipment removes impurities from coal, has been most difficult to evaluate in the past. The most widely used yardstick is the Frazer and Yancey efficiency formula developed in 1922,1 but Yancey in a later article states that "washers treating coals of different density composition or operating at different densities of separation cannot be compared directly on the basis of this criterion."' Prior to and since 1922, a variety of other methods has been used for comparison purposes, including the distribution curve, the error area, and the "ecart probable" or probable error. Yancey and Geer in discussing these methods conclude, "Performance can be evaluated in a number of different ways, with the choice of the proper method to use being dictated by the objectives of the investigation and the data available."' It is true that performance can be evaluated in a variety of ways, but if the equipment is to be evaluated on an effectiveness basis, there should be only one universal comparison method. Varying methods have been used because one universal comparison method has not been found or developed. In the article previously quoted, Yancey and Geer state in clear terms the primary concept for a universal comparison method: "One of the simplest, and certainly one of the most obvious evaluations of washery performance is the quantity of sink material in the washed coal and the float material in the refuse. If the washery products are tested at the density at which the washing unit is operated, the sink in the washed coal and the float in the refuse represent material that has been misplaced." The quantity of misplaced material was used as a criterion of washery performance by Lincoln in 1913,8 by the United States Bureau of Mines in 1938,4 by Hancock in 1947,3 and by the national French research agency Cerchar in recent years.2 In 1950 Anderson' proposed the use of this criterion as an efficiency value to replace the Frazer and Yancey formula. However, none of the above-mentioned investigators used the misplaced material concept in a manner that would provide universal coal-cleaning equipment comparisons. The Correlation Theory The ideal coal cleaning process would treat all sizes and would make a perfect separation at any given specific gravity. All material lower in density than the desired value would report in the coal product and all material higher in density would report in the refuse product. Unfortunately, no known cleaning process achieves this goal and there seems little likelihood that any process yet to be invented will do more than approach it. When coal is treated in volume under operating conditions, it is impossible to avoid mechanical entrapment, fluctuations in throughput and effective gravity of separation, and the creation of turbulent currents, even when a true heavy-liquid bath is used and the feed is closely sized and contains little intermediate gravity material. This being so, it is possible to appreciate the difficulties inherent in trying to obtain a perfect separation when treating a wide range of sizes and a feed containing high percentages of intermediate material, using turbulent currents to help create the effective separation gravity, under operating conditions which normally tend to be on the overload side. When coal is separated from refuse in any coal cleaning equipment, some refuse always reports to the coal and some coal to the refuse; the writer therefore assumed that there should be a relationship between the total amount of misplaced material produced by any given piece of equipment and the difficulty of separation as represented by the percentage of near gravity material in the feed. With small amounts of near gravity or ±0.1 material in the feed there should be less misplacement of material than would occur with large amounts of near

Jan 1, 1952

By E. Martinez, M. L. Hollander

Sulfurization of copper oxides was investigated using differential thermal analysis (DTA) and thermal gravimetric analysis (TGA). When mixtures of cuprous oxide and sulfur were heated, the surface of the cuprous oxide was converted to cupric sulfide at the melting point of sulfur (119°C); additional sulfuriza-tion took place when the sulfur vaporized above 200°C. Cupric oxide also reacted with sulfur to form cupric sulfide but the starting temperature was approximtely 225°C. When the cupric oxide and sulfur were physically separated, sulfurization still occurred indicating a reaction with the vapor phase of sulfur. Sulfurization was obtained in runs with cuprite, chrysocolla and basic copper carbonate. Further decomposition of cupric to cuprous sulfide was noted at somewhat higher temperatures. The DTA-TGA data made it possible to set up postulated reactions in each system. These techniques were found to be well suited for studies of this type. Because of the difficulty encountered in flotation of many metallic oxide minerals, efforts have been made to find suitable reagent combinations to sulfi-dize these minerals in the flotation circuit. Recently Bautista and Sollenberger,1 using a different approach, showed that the surfaces of many oxide minerals can be converted to sulfides by heating them with sulfur to temperatures between 150° and 225°C. This technique was effective even with chrysocolla, a mineral which is very difficult to sulfidize. The primary objective of the present study was to obtain information on the reactions between the oxide minerals and suIfur. A better understanding of the mechanisms involved might prove useful in pointing the way to improvements in the process. A secondary objective was to demonstrate the value of thermo-analysis in investigations of this type. Bollin and Kerr2 have reported the use of a modification of differential thermal analysis to study the thermal reactions which occur on heating elemental constituents in a closed system to synthesize copper sulfides and iron sulfides. Differential thermal analysis (DTA) was used to detect enthalpic changes as mixtures of minerals and sulfur were heated. The DTA indicated the temperatures at which reactions started and whether they were exothermic or endothermic. Weight change data were obtained by thermal gravimetric analysis (TGA). These data could often be used to support postulated reactions and to eliminate consideration of other possible reactions. EQUlPMENT Differential thermal analysis (DTA) involves measuring the temperature difference between the sample being tested and a thermally inert reference, generally aluminum oxide, as both are heated at a uniform rate. Series connected thermocouples in the sample and reference are used to detect the temperature differences that occur as the sample undergoes enthalpic changes.3 Crystalline transitions, fusions, vaporizations, dehydrations, decompositions and many solid state reactions can be detected. A DTA unit manufactured by the Robert L. Stone Co., Austin, Tex., was used in this investigation with a temperature rate increase of 10°C per min. The samples and reference were placed in Inconel cups with wells in the bottom into which the differential thermocouples fit, and run in a nitrogen atmosphere. Thermal gravimetric analysis (TGA) measures the changes in weight that a sample undergoes as it is heated. A Chevenard thermobalance converted electronically for graphic recording was used with a heating rate of 5°C per minute. An X-Y recorder plotted temperature on the X axis and weight changes on the Y axis. A flow of 250 ml per min of nitrogen was streamed through the furnace. Samples were placed in a porcelain crucible open at the top, unless otherwise noted. EXPERIMENTAL RESULTS Sulfur: Two sulfur samples from different sources were tested. The DTA of a high purity sulfur and a reagent grade sublimed sulfur are shown in Fig. 1. Nearly all the tests were run with the sublimed sulfur due to the difficulty in grinding and high electrostatic charge encountered with the high purity sample. Sulfur undergoes three endothermal reactions below 200°C. The first is a solid-solid crystalline transi-

Jan 1, 1964

By Maurice Lacasse, Pierre Legast

GENERAL OUTLINE For more than one hundred years, asbestos ore has been mined on Les Mines d’Amiante Bell's property at Thetford-Mines , Quebec, Canada. In 1878, the first tons of fibre were produced. The mine operated as an open pit until 1951 at a yearly production of about 330 000 mt of ore. Since 1951, the mine has operated underground using a block caving method. The mine produced 725 000 mt per year until 1978, when the rate of production was increased to 1 100 000 mt. The ore was produced entirely fm the first lift on 650 level until 1976. Then, a second lift on 950 level was put into operation to replace production facilities fm the first lift which is now exhausted. Annual fibre production is 72 500 mt with grades £ran 3 to 7. The total work force is 600 people. The host rock is a peridotite partially serpentinized. The asbestos fibre seams occur as stockwork deposit structurally controlled by talcose shear zone. The geological rock formations are heavily jointed and have a wide range of competence. Asbestos ore bodies have an heterogeneous nature which des difficult geological interpretation. THE GRIZZLY METHODS A full gravity method was used on the first level where the caved ore was drawn from finger raises onto grizzly bars spaced at 480 mn. Oversize blocks were reduced by secondary blasting. The ore fell by gravity in- to the ore pass below the grizzly chamber and was loaded into 5.7 mt haulage cars through a 1.2 m wide chute. Access to the undercut drift was fm the grizzly drift through finger raises. The undercut trough was drilled and blasted in a 1.2 m interval ring sequence into a slot situated at the end of each panel. The undercut stope was emptied through the finger raises. The undercut retreated from South to North at blast interval determined by the finger raises spacing. Fig. 1 represents the various steps involved with the grizzly method. Extraction drift dimensions were opened 1.8 m wide x 2.4 m high with anchor rock bolts and wire mesh where safety required them during excavation. Grizzly chambers were excavated at 2.4 x 2.4 x 4.6 m long on a 7.6 x 7.6 m grid spacing. Ore passes with dimensions of 1.5 x 1.5 m were developed from the haulage drift. On the extraction level, headings were drilled with jacklegs and cleaned by slusher into ore passes to the haulage chutes. The support used in the grizzly chambers was square concrete structure with 150 mn H beam reinforcement. The H beam was partially buried in the concrete and had one free face to allow for difference frequency of blast vibration between concrete and steel. The supported brow of the chamber had an opening 1.5 x 2.1 m. Ore passes were ring concreted 150 mn thick without any reinforcement. A 24 MPa compressive strength concrete was pumped and poured in pre-assembled forms. Based on the success of this method on the first level, the same layout was repeated on the second level. With the start of undercutting on the second lift, signs of deterioration were noticed in extraction drifts. At first, there was great concern as the operators were confident that standard repair techniques could cope with such a situation. An intensive program of support repair and a reduction of draw resulted. The following method of repairs were used: 1) Steel square sets of 150 mn H beam at 0.6 m intervals were installed. 2) This was reduced later to 0.3 m interval and eventually to a skin to skin steel structure. 3) Yieldable arches were introduced spaced at 0.9 m intervals, linked laterally with 150 m. H beam. 4) The yieldable arches spacing was then reduced to 0.46 m and eventually installed skin to skin. However, this support was not successful.

Jan 1, 1981

By W. A. Dickinson, H. F. Hendershot

THIS paper is a brief discussion of the past and present methods employed by the Huntington Works of The International Nickel Co., Inc., in the production of nickel and high nickel alloy seamless tubing. The various operations for both hot and cold processing will be considered in their natural sequence, together with the processing characteristics of the high nickel alloys and some descriptions of the equipment used. The tools and lubricants currently in use will be described. This discussion will relate primarily to nickel, Monel,† and Inconel,† since these are the most common of the many alloys processed into tubing at the Huntington Works. The nominal compositions of the materials are shown in Table I. CONVERSION FROM SOLID BILLETS TO TUBE SHELLS History The commercial production of seamless nickel and high nickel alloy tube shells is a comparatively recent development, the first nickel shells having been produced in 1922, Monel in 1930, and Inconel in 1931 on rotary-type piercing mills. Rolled or forged billets, ranging from 3 ½ - to 7 ½ in. diam, were pierced to produce shells ranging' in size from 4 3/8 in. OD X 0.180 in. wall to 8 3/8 in. OD X 1.0 in. wall. After piercing, the shells. were further processed by hot rolling, reeling, and sizing. Subsequent to these operations the shells were either supplied to the trade as "hot finished" tubes, or were processed by cold drawing to smaller sizes at the Huntington Works of the International Nickel Co. The extrusion of nickel and high nickel alloys also may be classed as a relative newcomer to the field of hot-forming methods, for it was not until 1928 that the commercial extrusion of these alloys was begun on a 600 long ton vertical extrusion press installed at Zenith Works of Henry Wiggin and Co., Ltd., Glasgow, Scotland. This press did considerable pioneer work and the results were so favorable that in 1937 a -2200 long ton horizontal press was installed. The quality of the inside surface of the tubing produced on this press was superior to that produced by piercing on a rotary-type mill and, as a result of this experience, a 4000 ton horizontal extrusion press was installed at the Huntington Works in 1943. Billet Preparation The first step in the extrusion, process is the selection of ingots with good hot malleability, as indicated by the test described by Martin and Bieber.1 Ingots then are forged into rough rounds which are machined to the desired billet diameter on a "bar peeler." This machine consists of a rotary head which carries the cutting tools and is driven by a 200 hp, dc motor. The feed mechanism consists of hydraulically driven input and output carriages on the respective sides of the cutter head. The diameter of the forgings used and the machined size or "billet diameter" that corresponds to the four different sizes of

Jan 1, 1951

By L. O. Bacon

IN the spring of 1952 Calumet and Hecla Inc. began a geophysical program near Shullsburg, Wis., in the Wisconsin-Illinois lead-zinc district, to assist the geological and drilling exploration programs. Sphalerite, the ore mineral in this area, occurs with a considerable amount of marcasite. In general the higher the percentage of sphalerite in the ore, the higher the percentage of marcasite; however, there may be mineralized zones high in marcasite but low in sphalerite. Orebodies in general are relatively narrow and elongated. Reconnaissance exploration of virgin areas consisted of drilling along lines at 500 to 1000-ft intervals. Whenever a mineralized zone was intersected additional drilling was done to determine whether or not the zone carried zinc percentages of ore grade and, if so, to delineate the body. Calumet and Hecla Inc. was interested particularly in a method that would show the direction and position in which to drill from the original discovery hole. This suggested a method of contacting the orebody, which occurs at depths ranging from 100 to 200 ft, with one electrode. After a number of procedures had been considered it was decided to use equipotential line methods. Instrumentation and Field Procedure: The power source used was a 400-cycle Signal Corps motor-generator having a rated output of 1200 w at 115 v. The detection unit consisted of test probes, a two-stage amplifier, and &apos;head phones for determining the equipotential line by audio null. Field procedure consisted of contacting the sulfide body in the hole by a specially built spring-loaded expanding electrode which would be lowered into the hole, in the closed. position, until it reached the mineralized zone. At that point the mechanism was released so that the arms would make contact with the walls of the hole in the same manner that a well-bore caliper does. This electrode was then connected to the generator, the other terminal of which was connected to the surface electrode. The surface electrode was either a point, a straight line, or a circular line electrode. The point surface electrode was placed at a considerable distance from the hole, that is, 2500 to 5000 ft, depending somewhat upon the depth to the sulfide zone and the distance from the hole containing the subsurface electrode to which exploration was considered. A higher voltage output from the power source would have been desirable, although no difficulty was encountered in determining null points to a distance of 500 ft. Straight line electrodes used were approximately 1000 ft in length at a distance of approximately 1000 ft from the hole containing the subsurface electrode. The circular line electrode used consisted of 4400 ft of bare No. 10 copper wire, which provided a circle of 700-ft radius. All line electrodes were fastened to the ground at 20-ft intervals with copper staples 4 in. long. Equipotential points were determined at 25 to 100-ft intervals. Location of the points was made by plane table and stadia. In open country with a three-man crew, the points were plotted as they were determined. With a two-man crew, pegs were placed at equal potential points and surveyed in after several equipotential lines were determined. Although interpretation of results given in this paper is qualitative, the paper as a whole would not be complete without a discussion of the theoretical distribution of potential due to a circular line electrode. The theoretical distribution would be required if a quantitative interpretation of the measured potential distribution were to be made. The theoretical potential distribution due to a circular line electrode can be obtained by means of Legendre polynomials. Using the method as developed by Weber,&apos; the potential for a circular line electrode in a semi-infinite, homogeneous, isotropic medium can be expressed, using the spherical coordinate system, as: V(r,n) = pi/2pR[S(r/R)nPn(cos?)Pn(cosa)]r<R [1] er V(r,a) =pI/2pr[S(r/R)Pn(cos?) Pn(cosa)]r>R [2] where: 0 is the angle at the origin, taken on the axis of the circular electrode between the axis and the line drawn to the circumference of the circular electrode. a is the angle between the axis and the radius, r.

Jan 1, 1957

By Carl Wagner

AT elevated temperatures. most metals react with oxygen, sulphur, or halogen rather rapidly, although a coherent layer of the reaction product is formed and separates the two reactants from each other. The reaction proceeds, since one of the reactants or both dissolve and diffuse across the reaction product. In many cases the rate is diffusion-controlled as is indicated by the validity of the parabolic rate law of Tammann&apos; and Pilling and Bed-worth. wott and Gurney&apos; have suggested that diffusion processes also play an important part in the case of reduction reactions. This is confirmed by observations reported below. In general, however, the rate of reduction reactions is determined not exclusively by diffusion processes. Reduction of Silver Sulphide Above 179°C silver reacts with liquid sulphur very rapidly. Since the volume of silver sulphide is much greater than that of the equivalent amount of silver, a coherent layer of silver sulphide is formed in accordance with a rule established by Pilling and Bedworth.&apos; According to Wagner&apos;,&apos;&apos; silver ions and electrons migrate from the Ag-Ag,S interface across the silver sulphide layer to the Ag,S-S interface where the following chemical reaction takes place: 2 Ag +2 e + S Ag,S If silver sulphide is reduced by means of hydrogen, the equivalent volume of the reaction product, silver, is less than that of the initial substance. Accordingly, no coherent layer of silver is formed. Instead, fine filaments of silver grow from the silver sulphide surface into the ambient gas. This phenomenon has been investigated by many authors, especially by Kohlschiitter," who has also reviewed previous literature. Kohlschiitter has interpreted the observed phenomena by migration of "silver" dissolved in silver sulphide. According to Wagner,&apos; silver ions and electrons are the migrating particles. The sequence of consecutive steps involved in the reduction of silver sulphide is shown in Fig. 1. Every- where at the silver sulphide surface, hydrogen reacts with sulphur so that excess silver ions and electrons are formed, which migrate at random until they are captured by a nucleus of metallic silver. If such a nucleus lies on the surface and material is furnished only from the base, the original nucleus is pushed outward and thus a filament grows. Removal of sulphur and formation of metallic silver may also take place at different points when silver sulphide is reduced by copper. The reaction is usually formulated as: AgL,S + 2 Cu = Cu.S + 2 Ag but the actual reaction products are silver and a solid solution of the two sulphides. Primarily, excess silver and electrons dissolved in silver sulphide are formed. After a certain degree of supersaturation has been reached, nuclei of metallic silver may form. If, however, a silver foil is in contact with the silver-sulphide phase, as is shown in Fig. 2, silver ions and electrons will migrate to the Ag,S-Ag interface where they are captured. To confirm this scheme, a composite sample, with the geometry shown in Fig. 2, was made by pressing silver sulphide tablets 0.6 cm in diam and 0.4 cm high together with a copper cylinder and a silver foil in crucibles of iron, which does not react appreciably with silver sulphide at 400°C. After heating for 3 to 7 hr at 400°C. the specimens were cut normal to the phase boundaries. It was found that the thickness of the silver foil had increased from 0.01 cm as initial value to a maximum value of 0.06 cm. In the interior of the silver sulphide tablets only traces of metallic silver were observed. These observations show clearly that silver ions and electrons may migrate over distances of several millimeters before they form metallic silver. Reduction of Cuprous Sulphide by Iron The behavior of cuprous sulphide is very similar to that of silver sulphide, since the lattices of the high temperature modifications are identical, but in Cu.S there is a variable deficit rather than an excess of metal. To investigate the displacement of copper by iron, cuprous sulphide powder covered by a copper foil was pressed in iron crucibles and heated up to 800 °C for 3 to 7 hr. The thickness of the copper foil increased from an initial value of 0.005 cm to a nlaximum value of 0.029 cm. In the interior of the

Jan 1, 1953

By J. C. McDonald

A DDITIONAL experimental data are presented in this note on a phenomenon which has been touched on but lightly in the literature. The common magnesium base rolling alloy (AZ31B) contains about 3 pet A1 and 1 pet Zn. When cold rolled, it hardens until it starts to break up in the rolls at 30 to 50 pet reduction. When it is heated after rolling to full hardness, it softens. However, other magnesium alloys have shown themselves capable of being cold rolled indefinitely, provided the ratio of roll diameter to starting gage is sufficiently high. As measured by the tensile test, these alloys actually soften when the reduction becomes large (60 to 90 pct). A subsequent heating at 300&apos; to 500°F can produce hardening of this metal. Heating at higher temperatures (600&apos; to 800°F) causes recrystalliza-tion and softening. Menzen&apos; has shown that a magnesium base alloy can be cold rolled without intermediate anneals as much as 90 pet. The alloy studied, Mg-1.5 pet Mn, softened considerably between 20 and 90 pet reduction, as measured by tensile properties; a decrease in yield strength of 30 pet, and an increase in elongation from 3.2 to 16.7 pet (longitudinal direction of the sheet). However, Menzen gives no data on the reaction of this metal. to heating. Ansel and Betterton2 and Hurst and Hatch3 report a similar effect in Mg-1.5 pet Mn-0.1 pet Ca. Their data show an initial hardening, followed by softening. The former authors also show that the alloy softened by cold rolling will harden when heated at low temperatures, and that these effects are absent in Mg-3 pet Al-1 pet Zn. In the course of work4, 5 some years ago, the author observed similar effects with alloys containing cerium, thorium, or zirconium; but not with elements like aluminum, silver, or zinc. Table IA contains the results of an experiment which shows these effects rather well. A binary alloy with 0.3 pet total rare earths (Mischmetal* added) was rolled hot to 1/16 in. and then annealed to full softness by heating 30 min at 800°F. Rolling was continued at 1.5 pet per pass to 80 pet total reduction on 8 in. diam rolls. The metal was harder after 20 pet work than initially but, at 80 pet reduction, properties were back to their original values. Heating for 1 hr at 450°F produced about the same hardness whether the reduction was 20 or 80 pet. At intermediate reductions, the metal hardened, but not to as high values. Extensive cold rolling may also be carried out on metal hot rolled at temperatures sufficiently low to produce a semihard temper. Mg-0.6 pet Zr was rolled to 0.1 in. at 700° to 800°F, 30 pet per pass. As shown in Table IB, a further cold reduction of 85 pet, at 1.5 pet per pass, definitely softened the metal.

Jan 1, 1959

By Robert W. Thomas

During the past 20 years the advantages of reinforced concrete as a substitute for timbering in so-called permanent mine openings have been fully recognized, and its use has become almost general practice in the larger mines. The term "permanent mine openings" is used advisedly and is assumed to embrace such openings as ventilation shafts, material shafts and main hoisting shafts, with their adjacent stations, ore pockets, etc. In the concreting of such openings, the actual cost as compared with that of timbering has been of secondary consideration, the support of large openings, elimination of fire hazards, and prevention of interruptions to production being the primary objectives. The ultimate economy, however, of lining such openings with concrete when they are to remain in use during the entire life of the mine, and especially when they are in bad ground, will seldom be questioned. During this 20-year period there have appeared in engineering publications a number of interesting articles on the use of concrete in mines, nearly always describing some specific project of major proportions, such as the concreting of ore pockets, large stations or main hoisting shafts. Such articles are unquestionably of interest, since the projects require considerable engineering thought, not only in the method of excavating large openings, but in their engineering design as well. Concreting of main hoisting shafts, particularly when this is done with the least possible interruption to production, provides some interesting problems. The concreting of haulage drifts of the dimensions commonly found in mines, on the other hand, soon develops into a routine operation, therefore it is with some hesitancy that the author attempts to prepare an article on this subject. The writer can recall only one article on the concreting of mine haulageways, although several articles have been written concerning the concreting of tunnels considerably larger in dimensions than those required in general mine practice.&apos; From the scarcity of articles on the subject, it would seem that the concreting of haulage drifts is still far from being considered general

Jan 1, 1937

By A. William Johnson

It was pointed out in an earlier paper1 that our understanding of the atomic mechanism by which diffusion occurs in metallic alloys is scarcely in an advanced state. This unsatisfactory condition is the result not of any lack of interest, but of the difficulty of designing experiments that .will yield suitable data. The diffusion coefficient of an alloy, measured in the usual way, is simply a number specifying the rate of homogenization of a nonhomo-geneous metal; this number, of itself, can furnish neither the rate at which the individual atoms move nor information as to the mechanism by which they move. It does not necessarily follow, however, that such information cannot be obtained, even if the path to be followed is still not clear. Much success has been attained in the study of diffusion in ionic crystals. A number of experiments, for which there are few counterparts in connection with metals, yield information often indicating a detailed mechanism by which diffusion probably occurs. From such work, quite acceptable theories of lattice imperfections have been developed by Wagner, Schottky, Frenkel, Jost and others.2 It is postulated that real crystals of stoichiometric composition contain, in equilibrium, certain imperfections of which the most importarlt types are: (I) vacant cation sites, with an equal number of interstitial ions; (2) vacant anion sites, with an equal number of interstitial ions; (3) equal numbers of anions and cations in interstitial positions; (4) equal numbers of vacant anion and cation sites. The classification of a particular crystal according to this scheme then follows from studies of the variation of electrical conductivity with pressure of the electronegative component surrounding the crystal, with ion size, and with temperature; other physical phenomena, such as the Hall effect, have aided in establishing the classification. The relative mobility of the anion and cation may be determined by chemical analysis of the salt at the electrodes following electrolysis. With such information, and knowing that one ion never occupies the lattice site normally belonging to an ion of opposite sign, it is usually possible to ascertain the diffusion mechanism. It is clear that the same experiments are not applicable to metal crystals, and substitute experiments that yield as much information have not yet been devised. Much can be learned, however, about the mobilities of the two* kinds of atoms in alloys by the use of radioactive metals as indicators. Only one such experiment has been performed in the past. In 1933, Seith and Keil,3 in an effort to explain the great difference between the rate of diffusion of gold in lead and of lead in lead, measured the rate of diffusion of radioactive lead in 3 gold-lead

Jan 1, 1942

By A. William Johnson

It was pointed out in an earlier paper1 that our understanding of the atomic mechanism by which diffusion occurs in metallic alloys is scarcely in an advanced state. This unsatisfactory condition is the result not of any lack of interest, but of the difficulty of designing experiments that .will yield suitable data. The diffusion coefficient of an alloy, measured in the usual way, is simply a number specifying the rate of homogenization of a nonhomo-geneous metal; this number, of itself, can furnish neither the rate at which the individual atoms move nor information as to the mechanism by which they move. It does not necessarily follow, however, that such information cannot be obtained, even if the path to be followed is still not clear. Much success has been attained in the study of diffusion in ionic crystals. A number of experiments, for which there are few counterparts in connection with metals, yield information often indicating a detailed mechanism by which diffusion probably occurs. From such work, quite acceptable theories of lattice imperfections have been developed by Wagner, Schottky, Frenkel, Jost and others.2 It is postulated that real crystals of stoichiometric composition contain, in equilibrium, certain imperfections of which the most importarlt types are: (I) vacant cation sites, with an equal number of interstitial ions; (2) vacant anion sites, with an equal number of interstitial ions; (3) equal numbers of anions and cations in interstitial positions; (4) equal numbers of vacant anion and cation sites. The classification of a particular crystal according to this scheme then follows from studies of the variation of electrical conductivity with pressure of the electronegative component surrounding the crystal, with ion size, and with temperature; other physical phenomena, such as the Hall effect, have aided in establishing the classification. The relative mobility of the anion and cation may be determined by chemical analysis of the salt at the electrodes following electrolysis. With such information, and knowing that one ion never occupies the lattice site normally belonging to an ion of opposite sign, it is usually possible to ascertain the diffusion mechanism. It is clear that the same experiments are not applicable to metal crystals, and substitute experiments that yield as much information have not yet been devised. Much can be learned, however, about the mobilities of the two* kinds of atoms in alloys by the use of radioactive metals as indicators. Only one such experiment has been performed in the past. In 1933, Seith and Keil,3 in an effort to explain the great difference between the rate of diffusion of gold in lead and of lead in lead, measured the rate of diffusion of radioactive lead in 3 gold-lead

Jan 1, 1942

By R. C. Ku, R. G. Davies

The stress-strain behavior of Fe-C and Fe-Ni solid-solution alloys has been investigated at 25°C. It is found that the rate of solid-solution strengthening is dependent upon the strain at which the yield stress is measured, the macroyield stress being much more concentration-dependent than the microyield stress. In fact the addition of nickel does not lead to any increase in the microyield stress of iron. The primary effect of increasing the carbon or nickel content is to increase the slope of the stress-strain curve in the microyield region; this effect has been correlated, by means of transmission electron microscopy, with a decreased ability of the dislocations to cross-slip. It is concluded that the Mott-Nabarro theory is a good description of true solid-solution hardening in these and other iron-base alloys. A.LL theories of the solid-solution strengthening1-4 indicate that the flow stress of an alloy increases both with solute concentration and "misfit" between the solvent and solute atoms. The functional relationship between strength, solute concentration, C, and misfit depends upon the particular theory. Mott and Nabarro1,2 and Schoeck and Seeger3 both predict a linear increase in strength with concentration, while Fleischer4 indicates that the strength should be proportional to C1/2; there are in the literature experimental observations to support both these compositional dependencies.1&apos;8 The experimental evidence used to support the theories is the change in macroscopic yield stress (strain, e, = 10-3) with composition. However, recent investigations have revealed that solid-solution strengthening in a Cu-Al single crystals7 and in Fe-V polycrystal-line" alloys depends upon the strain at which the flow stress is measured; a flow stress in the microplastic range (e = 10-6 to 10-5) is much less concentration-dependent than the macroyield stress. Since the theories of solid-solution strengthening are concerned in particular with the stress to move a dislocation in a dislocation-free matrix, they should only be compared with microyield-stress data; it will be shown that in a material deforming homogeneously the microyield stress (e = 10-6) is approximately the stress to move through the matrix those dislocations which have been nucleated by stress concentrations such as particles or grain boundary ledges.9-11 In contrast, the macro-yield stress is more complex in that it is affected by several phenomena: dislocation multiplication, dislocation interactions, and the ease of propagating slip across grain boundaries. For example, it is experimentally observed that when long-range order12&apos;13 or solute effects10 restrict cross alip the grain boundary

Jan 1, 1967

By J. B. Hess, R. L. Dietrich

In the mechanical twinning of magnesium on the {1012} planes the crystal-lographic deformation is such that, in the direction of the hexagonal axis [0001], twinning is possible only undef tension stress, while in the directions perpendicular to the hexagonal axis, twinning is possible only under applied compression. Thus in extruded magnesium and magnesium alloys where the extruded textures are best described as either [1010] or [1120] fibers,&apos; the basal planes are all nearly parallel to the extrusion axis and such material is susceptible to ( 1or2 ) twinning by compression in the direction of the extrusion axis. In order to explain the results of cyclic stressing in alternate tension and compression upon the tangent elastic modulus in such material, Dorn and Thomsen2 have postulated that, following twinning under applied compression, a retwinning or un-twinning occurs as a result of residual microtensile stresses immediately upon removal of the applied compressive force. Recently, Carapella and Shaw4 have used this supposition of untwinning in magnesium in discussing their results on indentation hardness and on cold drawing of MI sheet. In trying to confirm this hypothesis by suitable direct experiments, the authors have been unsuccessful in detecting any evidence whatsoever that untwinning follows the release of applied compression and are forced to conclude that this hypothesis is improbable or, at least, that the quantity of material involved is small. For the experiments, small cylindrical specimens 0.040-in. diam. X 0.120- to 0.140-in. long were etched from 0.0625-in. extruded rod. The ends of the specimens were accurately faced off parallel in a jeweler&apos;s lathe. The compression jig was simply an ordinary set of micrometer calipers with an anvil interposed between the spindle and the specimen. The anvil was free to move axially but restrained from rotating so that the rotation of the spindle would not introduce a twisting stress on the specimen. With this arrangement the compressive strains could be estimated to 0.0001 in. and X ray pinhole photographs could be taken with or without load as required. Fig I is the pattern of an uncompressed specimen of AZ3I alloy, showing the approximate [1010] extrusion fiber. The X ray beam is normal to the extrusion axis; in this position it is particularly sensitive in detecting twinning if one directs his attention to the diffraction ring from the basal plane. Since twinning on (1072 planes produced an 86" reorientation of the crystal lattice, the original basal reflections occur at the 3 and 9 o&apos;clock positions while the (0002) reflections from any twinned material will occur at the 6 and 12 o&apos;clock positions. Fig 2 shows the same specimen compressed 3.0 pct, photographed while still under load. This strain has been just enough to produce a detectible amount of twinning and the sensitivity is roughly indicated by the fact that no alterations were produced in the intensities or maxima of the other diffraction rings. A strain of

Jan 1, 1949

By R. L. Bleifuss

The object of this study was to evaluate the oxidized taconites of the western Mesabi iron range and to establish a correlation between the various basic taconite types and their concentratability. The initial work included a study of the original mineralogical and stratigraphic facies of the fresh taconite and traced the oxidation and leaching patterns developed during the conversion of this material to oxidized taconite. Subsequent studies have extended this work to include the semitaconites and also the wash and heavy media ore types widely developed in the area. Each basic original mineralogical facies or taconite type produces during alteration an equivalent secondary oxidized facies which has a unique mineralogical composition and texture. These control the concentratability of the oxidized taconites. Actually the influence of the original mineralogical composition of the iron formation extends even further and determines whether the most likely product of oxidation and leaching will be a wash ore or a heavy media ore and even to the direct shipping ores where the ratio of hematite to goethite is largely determined by the original taconite type. The relationship between original taconite, oxidized taconite, semitaconites, and wash and heavy media ores is briefly summarized in Table I. The terms are not precise and in some instances tend to overlap. The table is intended to point out only the essential differences between the various materials. The original taconite was dominantly composed of mixtures of magnetite, iron carbonate, iron silicate, and quartz (chert). In a few thin stratigraphic units, hematite was present as an original constituent. The alteration of the original taconite to oxidized taconite has resulted from the circulation of ground waters through an initial fracture system related to minor structures in the iron formation. The chemical reactions have involved the oxidation and decomposition of original minerals such as magnetite, iron carbonate, and iron silicates to hematite or goethite. The alteration at this early stage is characterized by pseudomorphic replacement of the original minerals by oxidation products although retaining the major textural features of the original taconite. There was very little secondary migration of iron at this stage and the total iron content of the rock has been only slightly increased. Further alteration of this material to form a semi-taconite, a wash or a heavy media ore, involves extensive removal of silica in solution and significant redistribution of iron units. Wash and heavy media ores are developed by the growth of coarse secondary iron layers, dominantly goethite, coupled with the removal of silica in solution and decomposition of the cherty layers. This paper is concerned primarily with the materials classified as oxidized taconites which are characterized by oxidation of the original mineral constituents with little secondary enrichment. The area studied, shown on the index map, Fig. 1, extends some 30 miles along the strike of the Biwabik formation roughly from Hibbing to Coleraine. This study was greatly facilitated by the basic stratigraphic and mineralogical studies of Gruner1 and white2. The bulk of the material in this presentation represents a portion of the author&apos;s contribution to a study sponsored by the Great Northern Railway on the oxidized taconites of the western Mesabi.3 METHOD OF APPROACH A total of 106 drill holes, representing some 30,000 ft of drilling, were examined and sampled. The drill holes were logged and some 1500 composite samples were prepared for chemical analysis and bench scale metallurgical tests. These samples were analyzed for total iron and ferrous iron; and magnetite iron was determined by a standard Davis tube test at minus 200 mesh. Standard reduction roast tests using 150-gram samples of minus 10 mesh material were made on these 1500 composite samples with a mixture of H2, CO2, N2, and H2O at a temperature of 650°C. This roasted material was pulverized through 150 mesh and concentrated in the Davis tube. The iron recoveries and concentrate grades obtained on the roasted products were used to assess the concentratability of the oxidized taconites. Detailed mineralogical work was done on thin sec-

Jan 1, 1964

By J. T. Norton, Pekka Rautala

The phases and equilibria in the W-Co-C system have been studied by X-ray diffraction methods, metallographic technique, and thermal analysis. In addition to the 7 phase, two double carbides, called 8 and have been revealed. The compositions correspond to CO3 W6C2 and Co3-W10C4. The reactions leading to these phases have been explained and tentative diagrams of stable and metastable equilibria proposed. The basic reactions in sintering cobalt cemented tungsten carbides are discussed. THE W-Co-C system is of fundamental importance in practical carbide manufacturing as well as for the understanding of the sintering mechanism. Surprisingly little is known about this system, for the probable reason that all the important alloys are of two-phase structure and that the diagram Co-WC has been treated as a quasi-binary. This obviously is incorrect, because WC decomposes before melting. One ternary phase, 7, of composition Co3W3C has long been known. It was first studied by Adelskold, Sundelin, and Westgren,1 although the isomorphous iron-tungsten carbide was known earlier. There have been in the literature several reports of two 7 phases.&apos; " Also the 7 phase has been considered unstable by Takeda4 and Westgren.1 The Co-WC diagram has been studied by Wyman and Kelley5 and a quasi-binary diagram has been published by Sandford and Trent.2 Takeda has published a tentative Co-W-C diagram, considering both metastable and stable equilibria. However, the lines of two-fold saturation are shown only partially and it seems impossible to complete the diagram without violating the phase theory. Therefore it seemed desirable to examine the system in more detail. Experimental Procedure The alloys used in the present investigation were made of powders of tungsten, tungsten monocarbide, cobalt, and carbon and were of the grade used in manufacture of commercial cemented carbides. The powders were ground and mixed in small stainless steel ball mills, using balls of the same material. Benzene was used as a dispersing agent. The mixing period was 1 hr, since this was shown to give suffi- ciently good mixing of the powders without too great a contamination from the mill. After ball milling, the specimens were pressed in cylindrical or rectangular dies. No paraffin or other lubricant was used and the small compacts had sufficient green strength to be handled without difficulty. Several sintering furnaces were employed. The most satisfactory arrangement was a vacuum furnace based on the Arsem principle which employed a graphite helix as the resistance heating element. Specimens were placed on graphite stands and there was generally a slight carburization or decarburiza-tion of the specimen surface, depending on the carbon content of the alloy. The evaporation of the cobalt at a sintering temperature of 1400°C was not significant, but became severe at 1500°C and higher. The sintering time was 1 hr at 2000°C and 2 to 4 hr at lower temperatures. It was not possible to quench the specimens, but the cooling rates were rather fast, greater than 300°C in the first minute. In the system under investigation, the reactions are sluggish, and it is believed that the high temperature structures are satisfactorily retained. The principal method of investigating the sintered specimens was X-ray diffraction by the Norelco recording spectrometer. Approximate determinations of the phase boundaries were made by the disappearing phase method. Ternary Phases To study the phase formation in W-Co-C system, a series of specimens was sintered at 1400°C. In this experiment two ternary phases, called here ? and k were formed in addition to the well-known 7 phase. The 7 phase, which has been completely described by Westgren: showed a range of homogeneity from 7 to 20 pct C and from 38 to 48 pct Co. At 1400°C the 7 phase was found to be in equilibrium with monotungsten carbide, 8, tungsten, 8, #?, and liquid. The boundaries toward #? and liquid were difficult to determine and appeared to be very temperature sensitive. The other boundaries are believed to be well fixed. The homogeneity range of 7, as measured in this work, is considerably smaller than the one

Jan 1, 1953

By Abram S. Hewitt

GENTLEMEN : If my first words were other than those of thanks for the high honor of being called to preside over the American Institute of Mining Engineers, I should do injustice alike to you and to my own sense of the obligation, under which I have been placed by your voluntary choice. When the position was first suggested to me, I resolutely declined to allow the use of my name, on the ground that, by training and occupation, I was not entitled to this honor, which, it then seemed to me, should be conferred only on a professional engineer. In fact, the history of the Institute, which had honored itself, as well as my distinguished predecessors, David Thomas, Rossiter W. Raymond, and Alexander L. Holley, by elevating them to its highest office, seemed to indicate the line and the limit of safe precedent, in which I did not then see how I could be included. But my scruples were finally removed by the consideration that, as in the course of human industry, the pioneer, of whom Mr. Thomas was so marked a type, must precede the mining engineer, so fitly personified in Dr. Raymond, and be followed by the mechanical engineer, whose incarnation we behold in Mr. Holley; so all these must be supplemented by the man of affairs and finance, in order that "enterprises of great pith and moment" may be prosecuted to a successful issue. I felt, therefore, that I had no longer any right to resist your preference, or to deprive myself of what I must regard as the crowning honor of an active career. On an occasion like the present it seems appropriate to review the history in this country, during the last hundred years, of those industries which the Institute of Mining Engineers specially represents. There is much in the story with which other members of this body are more familiarly acquainted than myself; yet I fancy that to all of them, and not less to our distinguished and most heartily welcome guests from abroad, a comprehensive outline of the progress of mining and metallurgy in the United States will prove interesting. I ask your attention, therefore, to a condensed account of that subject, embracing a statement of the beginnings of the industry on our shores, of the notable events attending its increase, and of its present

Jan 1, 1877