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By W. C. Coffin

The refractory linings of open-hearth steel furnaces above the bath line are subject to severe wear not only from the heat caused by the combustion of the fuel and the reactions of the bath, but also from the melting action of the gases when not controlled and directed, and from the scouring of the oxides, etc. with which the products of combustion are charged. The framework that encases or binds the furnace is made up primarily of an iron or steel pan in which the furnace proper is built, side and end buckstays, and tie-rods. The expansion and contraction are so great and variable that the bindings may be subjected to severe stresses unless the tie-rods are carefully adjusted in heating up the furnace, but they are most affected by direct heat and usually become deformed from that cause when the lining is worn thin and, sometimes, when sections of the lining fall in and expose some of the framework. A deformed framework displaces the skewback of the roof arch and tends to distort or crush the roof. Bent buckstays deform the front wall, causing the door frames to stand away from the lining so that the gases burn between the lining and the frames, destroying both of them. Frequently the frames are attached to the buckstays; then when the latter bend the frames usually break. Preventing the distortion of the framework and maintaining the economy of the refractories are real factors in the cost of steel making. Nat only does the cost of general repairs and reconstruction add directly to the cost of steel produced, but the time lost in making repairs cuts down the steel production and increases the overhead expense of both the furnace shop and its allied foundry or mills. Many years ago water-cooled equipment applied to iron blast furnaces made it possible to increase greatly not only the size and production of the furnaces but also the life of the lining. It also cut down the loss from frequent repairs. The open-hearth furnace is a much later invention, especially as a commercial factor; for though it dates from the Fred Siemens patent of 1856, much more progress was made in the Bessemer process until the latter eighties and most of the development

Jan 1, 1920

By AIME AIME

MORTIMER ELWYN COOLEY, dean of the College of Engineering and Architecture of the University of Michigan, has been elected president of the American Engineering Council of the Federated American Engineering Societies. This was announced at a meeting of the Executive Board of the Council held at the Cosmos Club in Washington on Sept. 30. Dean Cooley assumes office at once. The new head of organized engineering is a past president of the American Society of Mechanical Engineers and has a long record of distinguished service in education, under the government and in private capacities. He was born in Canandaigua, N. Y., March 28, 1855, and was graduated from the United States Naval Academy in 1878. With his classmate, Dr. Ira N. Hollis, he was ordered to the U.S.S. "Quinnebaug," then fitting out at the Norfolk Navy Yard preparatory to a cruise in the Mediterranean.

Jan 1, 1921

By C. Dinis da Gama

In order to reduce the complexity of mechanisms influencing rock. fragmentation by blasting a simulation approach is proposed, using the capabilities of microcomputer interactive graphics. Situations involving cratering end bench blasting are simulated, and both jointed and intact rocks can be subjected to explosive action, taking into consideration its breakage mechanisms. Size analysis curves relating rock distribution before and after the blasts are determined and compared with field data.

Jan 1, 1984

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Horace V. Winchell

The St. Helens mining-district, indicated in sketch-map, Fig. 1, is chiefly in Townships 9 and 10 North, Ranges 5 and 6 East of the Willamette meridian, in Skamania county, Wash. There is no railroad or even a wagon-road at present connecting it with any base of supplies; access is usually by wagon-road for about 50 miles from the town of Castle Rock to Spirit lake, and from that point by pony-trail from 6 to 10 miles farther. History. Nearly 20 years ago two German farmers, totally inexperienced in prospecting, penetrated this region upon a hunting-and fishing-trip. Their attention was drawn to numerous joints and crevices filled with black, glistening tourmaline, and con-

Jan 1, 1913

By C. N. Kravig

STANDARD raises at Homestake are 150 ft long. Raises of more than 150 ft, driven with the use of a cage and hoist, are considered long raises- some have been driven more than 900 ft. Long raises have eliminated the need for access drifts on each level when transfer and ventilation raises are driven. These timbered cage raises, 7x9 ft in cross section, include a rock compartment and a combination cage, pipe, and manway compartment. To prevent excessive weight on the timbers, 10x10-in. bearing beams are cemented into hitches cut in the rock at 150-ft intervals. The beams also hold back timber movement caused by daily drawing of rock through the rock compartment.

Jan 6, 1957

By Aloys H. Wohlrab

[THE conditions that govern the selection of a suitable type of detachable bit for the small isolated mine, for rock work and tunnel contracting and for the large mine are quite dissimilar, therefore intelligent recommendation of a particular make of detachable bit must involve consideration of many factors. Small, isolated mines may find advantageous a detachable bit that a larger mine cannot afford to use, because of more efficient conventional steel sharpening and steel-distribution practice, but a small mine may of be able to adopt the practice of larger mines because of necessary capital expendites for equipment for detachable-bit sharpening and shanking. The perfect detachable bit, suitable for all conditions, has n yet been developed. CONVENTIONAL DRILL BIT IN USE That drilling speed increases as the hit gauge decreases is well known, but since small-hole drilling is the exception rather than the rule the rate of such increase is probably not so generally recognized. Tests married out by Holman Brothers Limited1 with 7/8-in. show that when bit sizes are reduced from. 1 ½ in. 1 3/8 in., and 1 3/16 in., to 1 3/8 in., 1¼ in., and it' in., respectively, the volume of the hole is15 per cent less and the drilling speed 4q per cent higher. Recent tests by the United States Bureau of Mines2,3in hard, uniform basalt showed that a reduction in average gauge from 2.077 to 1.826 in., a difference of 0.251 in., gave an increased drilling speed of 65 per cent. The limits to which a drill hole may be extended depend upon the loss in gauge. Since the conventional bit with 5° and 14° clearances did not possess sufficient reaming edge to permit the smaller gauge changes that were required for small-hole drilling, experiments were made in bit construction with a view to correcting this deficiency. This resulted in the adoption of the full reaming cross bit (having full Carr characteristics) by some mines.4 In order to drill holes having a minimum taper from the collar to the bottom of the hole, full reaming center-hole cross bits with 90° cutting angles, 5° and I4° wing tapers, forged on 7/8-in. quarter-octagon drill rods, are used in the conventional drill-steel practice at Hollinger. In the steel shop close attention is given to the condition of the dies and dollies, which, combined with a rigid inspection of the finished bit, ensures as nearly a perfect bit as can be forged economically. Shop practices and costs have been discussed in earlier papers.5,6 The method by which the radii for the special dies used for forging bits of different gauge diameter are obtained so that the points and reaming edges cut circles of the same diameter has been discussed by Hibbert.7 REQUIREMENTS FOR DETACHABLE BIT FOR HOLLINGER Under these circumstances, it was apparent that it would be difficult to introduce a detachable bit that could successfully compete with the conventional bit already in]

Jan 1, 1942

The Hadsel Mill. By R. G HALL (Trans, vol 112 4000 words) This mill is designed to effect the entire reduction of ores by combined crushing and grinding from mine run size to material sufficiently fine for the usual operation of flotation or cyaniding The work is begun and completed in the same machine. The machine is designed on the principal of the elevator wheel, and crushing and grinding are accomplished by dropping a mixture of ore and water upon specially placed "breaker plates" in such a manner that the resultant material passes back into the buckets of the wheel for further elevation One method of removal of the finished product is to permit an overflow of wet pulp on one side of the mill through coarse screens A recent improvement provides for the removal of a pulp by a launder part way up the path of the buckets In either case the pulp passes to a classifier, and the oversize back to the mill in the manner familiar in ball-mill practice Although hardly through the development stage, the machine has given a very good account of itself in simplification of plant, by elimination of coarse and fine crushers with consequent reduction of installation cost and power consumption As compared with the ball mill alone, there is a reduction of power consumption as well as a material reduction of grinding steel There is an increase in cost of maintenance other than grinding steel, but as the machine passes out of the development stage there is a show¬ing of reduction of costs on this item It seems probable that with further develop¬ment and strengthening of the weaker parts further improvement may be made Ball Milling. By A. M Gow, M GUGGENHEIM, A. B. CAMPBELL AND W. H COGHILL. (Tech Pub 517; also Trans, vol 112. 24,000 words) This paper outlines a new philosophy in ball milling by using power as an index of mill operation and ascertaining the effects of the set and induced variables By recognizing the amount of ore in the mill as an important induced variable and recognizing that the amount of ore in the mill depends principally on the size of the discharge trunnion, ball volume, rate of feed, and speed of mill, obstacles formerly met were surmounted When the amount of ore in the mill is dealt with intelligently and the applied work is proportional to the amount of feed, series testing, which would otherwise give a hotchpotch, gives concordant results The study of ball paths and trajectories was seen to be fruitless when it was shown that the old attrition theory for slow mill speed

Jan 1, 1935

By W. R. Ingalls

Jan 1, 1922

By W. R. Ingalls

Jan 1, 1922

THE developments in construction and operation of the basic open-hearth furnace that have taken place in recent years have placed new emphasis on the vital influence played by regulation and control of the combustion system in both furnace production and furnace life. The several factors influencing flame development, such as regulation of fuel supply, air supply, air preheat, furnace draft, and furnace reversal, must be controlled closely if successful furnace performance is to be accomplished. Rugged instrumentation of distinctive design is required to supply dependable information concerning operating conditions, which is requisite to proper regulation of combustion. This chapter attempts to give a practical discussion of the various features of open-hearth practice and instrumentation related to the efficient heating of the furnace, control of temperature, and protection of the furnace structure from accidental overheating. The appended references will be helpful to readers wanting a more detailed treatment of individual topics. OPEN-HEARTH FUELS The most important open-hearth fuels are listed in Table 4-1 together with their compositions and other data useful for combustion calculations. A detailed discussion of open-hearth fuels is not attempted here, but certain factors peculiar to their application to this type of process are briefly indicated on the following pages. Very often an open-hearth shop is forced by economic factors to use a fuel to which its furnaces are not very well adapted, because the general fuel-allocation scheme for the works as a whole seems to demand it. Often a variable marginal supply of coke-oven gas or tar must be consumed by the open hearth, together with another fuel such as producer gas, natural gas, or oil, to fill out total requirements. Most furnaces employ more than one kind of fuel during their usefu1 life; however, the partial redesign

Jan 1, 1951

By R. W. Raymond

FOUR portraits have recently been hung in the rooms of the Institute, in recognition of four illustrious men with whom we, as mining engineers and metallurgists, may claim fellowship. LUTHER. Martin Luther is often spoken of as a miner's son; but it is less generally known that John Luther, his father, was, in a small way, a metallurgist as well as a miner, and that Martin Luther himself, in his youth, worked in a forge or bloomary established by his father at Mansfeld, Saxony. Since this is the only manual occupation on his part of which we have any record, it is fair to assume that the great scholar and reformer derived from his work in this department the practical familiarity with the conditions of human labor so abundantly evinced in the homely illustrations which accompany even his serious arguments, and in the witty apothegms of his famous "table-talk." At the same time, I must confess that 1 have not been able to find in any of the works of Luther distinct allusions to mining or metallurgy. KÖRNER. Theodor Körner, a student at the Mining Academy of Freiberg, Saxony, threw himself into the German War of Freedom in the first decade of the 19th Century. He was a poet of no mean rank ; and his " Song of the Sword " is one of the great lyrics of literature. This poem was written but a few days before he died in battle. In many respects, Körner may be compared with our American hero Theodore Winthrop, who similarly met a glorious fate, and left behind him prophetic indications of genius, of which his premature death prevented the complete fulfillment. I am informed by a member of the Institute that Körner's admirable collection of minerals is still preserved at Dresden.

Nov 1, 1908

By W. C. Elmore

THE magnetic powder method, long used for roughly mapping mag-netic fields, has recently been refined 1,2 for investigating the microscopic variations in the surface magnetization of ferromagnetic crystals. Bitter2 first discovered regular powder patterns connected with the crystal structure of such specimens. Several other investigators3 have used Bitter's original technique, with slight modifications, to study the varia-tion of the patterns with variations in the magnetizing field applied parallel to the surface, and their dependence upon the state of strain, of polish and of crystal cut. All of the explanations of these patterns on metallic ferromagnetics have been couched in general terms. In a preliminary report we have described an improvement in the powder technique which led us to the discovery of a new effect throwing con-siderable light on the magnetic secondary structure of alpha iron4 and of nickel5 single crystals. We now report additional evidence confirming the existence and stability of a magnetic block secondary structure.

Jan 1, 1936

By F. Weinberg, E. H. McLaren

Dilute binary alloys have been solidified under controlled thermal conditions, and solute distributions, temperatures during freezing and melting, and the position and morphology of the solid-liquid interfaces have been examined. Binary Zn, Sn, and Pb based alloys were investigated with solute additions of AgU°, AU°, Sb'24, TlZo4, snH3, and Zns5. Approximately 150 cc of metal was used in each ingot. Appreciable solute segregation occurred in most of the ingots, depending on the solute concentrations, distribution coefficients. and freezing conditions of the alloys. Measurements were made of the general solute distribution by sectioning layers of material from the ingot circumference and counting the activity of each layer. Local segregation was detected by autoradiographic techniques. Freezing and melting curves for the same alloys were obtained by precision resistance thermometry. These curves are related to the corresponding measured solute distribution curves and those calculated from theoretical expressions. The position and morphology of solid-liquid interfaces during freezing were determined by both decanting and quenching techniques. The results indicate the manner in which freezing and remelting occurred under the conditions investigated. If an isomorphous dilute binary alloy is solidified or melted under equilibrium conditions, the concentration of solute in both solid and liquid would be uniform at all stages in the freezing process. The solute concentrations during freezing could be determined from the phase diagram of the alloy, and the freezing curves would have an alloy slope defined by the solidus and liquidus points. In practice, however, solidification under equilibrium conditions rarely, if ever, occurs, primarily because of low rates of diffusion in the solid and incomplete mixing in the liquid. As a result, the solute distribution and the shape of the freezing and melting curves obtained in a given system depends on a number of factors. These include, in addition to the extent of diffusion and mixing, the initial solute distribution and concentration, the distribution coefficient, the rate of freezing or melting, supercooling of the liquid, the morphology of the solid-liquid interface, and the effect of annealing. The purpose of the present investigation was to determine, experimentally, actual solute distributions and thermal curves for a series of dilute binary alloys, using tracer techniques to determine the solute distributions and precision resistance thermometry for the temperature measurements. In addition, observations were made of the positions and morphologies of the solid-liquid interfaces during freezing and remelting. The distribution of solute in a dilute alloy, assuming complete mixing in the liquid, and negligible diffusion in the solid, is given by1 where Co is the mean solute concentration of the alloy, C, is the solute concentration in the solidg is the fraction of the ingot solidified, and ko is the equilibrium solute distribution coefficient (assumed to be independent of concentration for dilute alloys). For the case of partial mixing, ko can be replaced by an effective distribution coefficient k. If solute transport in the liquid is assumed to be entirely due to diffusion, with negligible diffusion in the solid, then the solute distribution in the solid is given by Tiller et a1.' as where R is the rate of advance of the solid-liquid interface, D is the solute diffusion coefficient in the liquid, and x is the position of the interface from the start of freezing. The temperature at the interface TI is related to the solute concentration of the solid at the interface C, by the expression where Tm is the freezing temperature of the pure solvent and ml is the slope of the liquidus for the binary alloy being considered. Accordingly, by relating TL to the actual temperatures measured in the present experiments, solute distributions can be deduced from the temperature measurements. Using the above expressions, the measured solute distribution of a given alloy may be compared to the theoretical distributions, as well as to the distribution calculated from the observed freezing curve. EXPERIMENTAL The solvent materials used in making up the binary alloys were C. P. New Jersey zinc (99.999 pct), Vulcan Extra Pure Tin (99.999 pet), and Tadanac lead (99.998 pct). The solute elements were of 99.99 pct purity or better, and were activated in the

Jan 1, 1963

By A. S. Adams

A GLANCE at the list of papers1 that have been published since 1920 on the general subject of flotation suggests the variety of ideas that exist regarding the underlying cause of the phenomenon. Among these ideas, we find several coherent theories, but none that adequately and completely explain the phenomenon. Ralston2 and Bains3 suggested an electrical theory which has to do with the attachment of mineral particles to a bubble by the attraction of unlike electrical charges. This theory has its root in the fact that a bubble of air in pure water is electrified in such a way that the surface is negatively charged while the body of the water is positively charged. This result, which was first discussed by Helmholtz, also obtains when the water contains anything except a high concentration of hydrogen ions, or ions which are trivalent or tetravalent. The theory also holds that the mineral particles have positive charges on their surfaces, and so there is attachment between bubble surface and mineral particles by the old law of electrostatics, which states that unlike charges attract.

Jan 1, 1928

By J. Murray Riddell, George A. Morrison

This paper is a companion to the one by George A. Morrison on Mining a Deep Limestone Mine in Ohio.‡ Barberton is 8 miles west of Akron, Ohio, and 23 miles south of Cleveland. The underground mining of limestone at Bar-berton constitutes the world's deepest limestone mine. The stratigraphy is as follows: 0-20 ft. Loam, clay, sand and gravel. Water bearing. 2-40 ft. Fractured sandstone. Water bearing. 40-110 ft. Sandstone with interbedded shale, slightly fractured. Water bearing. 110-314 ft. Sandstone with interbedded shale, slightly fractured. Water carrying. 314-2197 ft. Shale with occasional seams of sandstone, which varies in thickness from a few inches to 4 feet. Dry. 2197-2247 ft. Limestone. Dry. 2247-2742 ft. Dolomitic limestone. Dry. 2742-2767 ft. Salt. Dry. 2767- Dolomitic limestone. Dry. All of the strata dip to the southeast, about 0.5 per cent from the horizontal. The top 110 ft. of the crust is water-bearing. The rock strata are naturally fractured to a depth of 314 ft., sufficient to allow the passage of water. In the shaft-sinking operations, a stra- turn of oil-bearing sandstone one foot thick was encountered at a depth of 520 ft., and flows of methane gas were pierced at depths of 1350 and 1952 feet. Specifications for the mine plant called for, in part, the sinking of two vertical shafts and a hoisting plant capable of producing at the rate of 300 tons per hour. No. I shaft is 550 ft. west of No. 2 shaft. Each shaft measures 8 by 17 ft. outside of steel. General features of the design (Fig. I) are: Principal set members, 6-in. H-section; connecting angles (studdles), ? by 4 by 4-in. angle irons; spacing of sets. 8 ft. centers; spacing of bearer sets, generally, 160 ft.; bearers, of 12-in. I-beams. Each bearing set featured a concrete ring about the periphery of the shaft, steel to rock, and 8 ft. in height; permanent guides, 6 by 10-in. long-leaf yellow pine, 85 per cent heart. The Shaft-sinking Operation The preliminary shaft-sinking work consisted of driving interlocked steel piling about the exterior of the shafts to depths of 35 and 32 ft., respectively, and of such horizontal dimensions as to ultimately permit the pouring of a concrete collar 2 ft. in horizontal thickness. Following the pouring of a shaft collar, a temporary wooden headframe for shaft-sinking purposes was erected. The head-frame, while ordinary, featured: (1) a storage pocket of 300 cu. ft, capacity with truck servicing facilities, and (2) pneu matically operated dump doors, which mere actuated by the hoistman, thus caring

Jan 1, 1946

By J. W. Cuthbertson, R. A. Cresswell

THE constitution of Cu-rich alloys with 1.5 to 13.5 pct Sn and 0.25 to 3.0 pct Be and the precipitation-hardening characteristics of alloys with 1.5 to 13.5 pct Sn and 0.25 to 1.0 pct Be have been examined. The hardness and tensile strength of the alloys examined increase markedly after solution treatment at 700°C followed by heat treatment at temperatures between 200" and 450°C. By a combination of cold work and heat treatment, hardness values similar to those exhibited by commercial Be-Cu alloys containing 2.25 pct Be can be obtained with ternary alloys containing 9 pct Sn and 0.75 pct Be and containing 10 pct Sn and 0.5 pct Be. Marked hardening effects occur with alloys containing even less beryllium. By heat treatment alone, a hardness value of 310 diamond pyramid hardness can be obtained from an alloy containing 10 pct Sn and 0.75 pct Be. Preliminary tensile tests have shown that an ultimate tensile strength of 110,000 psi with an elongation of 23 pct is obtainable by precipitation hardening an alloy with 8 pct Sn and 0.75 pct Be. The precipitation-hardening process has been followed microscopically for certain alloys and the inference is that, while the initial hardening effect is probably explained by the precipitation of the ß phase of the Cu-Be system, further hardening, proceeding at a much slower rate, also occurs, apparently as a result of precipitation of phases of the Cu-Sn system, particularly precipitation of the 6 phase at temperatures below 350". The presence of the e phase of the Cu-Sn system in certain alloys at temperatures below 350°C has been confirmed. Tin-bronzes are widely used in engineering applications where a combination of high strength and good resistance to corrosion is wanted. The maximum strength is induced in these alloys by cold working, and it would be an advantage for many purposes if high strength could be achieved alternatively by an age-hardening process. While Cu-Sn alloys have a good fatigue resistance they can be surpassed in this respect by Cu-Be, but the use of the latter alloy is limited by its high cost. If, by adding beryllium to tin-bronze, the properties of the respective binary alloys could to some extent be combined, a most attractive alloy should result. As pointed out by Raynor,¹ beryllium is on the borderline of the zone of favorable size factors for copper, and the solid solubility of beryllium in copper is consequently much more restricted than if the size factor were strongly favorable. The size factor is sufficiently favorable, however, to permit an increase in solid solubility with rise in temperature, and there is thus a composition range in which CU- Be alloys are susceptible to hardening by precipitation heat treatment. Although the a phase of the Cu-Sn system is similarly susceptible to precipitation treatment, the time necessary to establish equilibrium in commercial alloys of this type is usually so great that age hardening becomes impracticable. The addition of beryllium to Cu-Sn alloys would appear to offer a means of conferring on the latter useful age-hardening properties. Masing and Dahl² and others have, in fact, shown that the addition of beryllium to Cu-Sn a solid solutions renders these alloys susceptible to precipitation hardening and after such hardening confers on them an encouraging improvement in physical properties. If this improvement could be achieved by the addition of substantially smaller amounts of beryllium than are customarily found in binary Cu-Be alloys, the ternary alloys should possess economic advantages which might make them more attractive than the binary alloy for some applications. Binary Systems Copper-Tin: The constitution of these alloys is now reasonably well known and is summarized in the equilibrium diagram published by Raynor.³ The following observations, due to Raynor,¹ on the structure of those phases of the Cu-Sn system that are likely to be found in the ternary alloy system will facilitate the subsequent discussion on the examination of that system. The ß phase is an electron compound at the electron-atom ratio 3:2 and has a body-centered cubic crystal structure. This phase is stable only down to 586°C, at which temperature it decomposes eutectoidally into the a and y phases. The y phase has a structure that is also based on the cubic system. This phase is stable down to 520°C, at which temperature it decomposes eutectoidally into the a and d phases. The d phase is an electron compound (Cu³¹Sn8) which has a crystal structure analogous to that of 7 brass. This phase is stable from 590" to 350°C; on prolonged annealing at the latter temperature it breaks down into a mixture of the a and E phases. The e phase is an electron compound (Cu³Sn) having the electron-atom ratio 7:4. Its structure may be regarded as a superlattice based on the close-packed hexagonal system. This phase is stable from 676°C to room temperature. The primary solid solubility of tin in copper increases to a maximum of 15.8 pct as the temperature falls from that of the peritectic reaction to 586°C. The solid solubility remains constant from 586" to 520°C. At lower temperatures the solubility decreases progressively. Below 350°C the fall in solubility is pronounced and is associated with the precipitation of the e phase. This precipitation is very sluggish and does not normally occur under service conditions. Copper-Beryllium: The Cu-Be system has been investigated by Borchers' and others. Raynor5 summarized the present state of information on it.

Jan 1, 1952

By A. M. Tweedy, Felix A. Vogel

Since the publication of Prof. J. W. Richards's paper on The Schumacher Briquetting Process,' this process has been in operation on a practical scale in two plants in the United States, and a few notes on the results obtained may be of interest. An experimental plant, consisting of a toggle press and a hydraulic press, was installed at Johnstown, Pa., together with the necessary bins and other equipment, and an attempt was made to briquette a mixture of Mahoning fines and blast-furnace flue dust, in a ratio of 7 parts of ore to 3 parts of dust. The results were unsatisfactory, in that the briquettes made were not hard enough to stand the severe handling to which they were subjected in the various transferences necessary between the press and the furnace. The trouble was due to several causes: First, the mechanical arrangement of the plant, which did not lend itself to a proper hardening and loading of the briquettes into the cars. Second, the irrrgular composition of the flue dust, as to both its lime and its fine coke contents. Third, the variability of the moisture content of the ore, and its viscosity when wet. Fourth, the ratio of fine ore to flue dust; that is, it was attempted to bind in too large a quantity of fine ore for the small amount of flue dust used. The first trouble could be overcome, and has been improved upon in the plant as it stands to-day. The second trouble was important, because of the small amount of dust used as a binder, and might have been easily removed by mixing the flue dust of the different furnaces, which, however, was never done. The third difficulty was more serious, but could have been partly solved, had it been possible to meet this along with the fourth difficulty, by decreasing the percentage of ore in the briquettes. It was proposed to do this, but it seems that the problem at the works in question is, primarily, to prepare Mahoning fines for reduction in the blast furnace, and only secondarily to use up the flue dust, and a solution along this line did not seem to be interesting.

Jan 1, 1914

By H. D. Dahl, P. C. Thakur

This paper presents an analysis of a new tool for improved productivity in coal mining: the horizontal drilling system developed by Conoco/Consol. The analysis includes a brief description of the system; discussion of the three categories of advance exploration application of horizontal boreholes; and a survey of other special mining methods planned around long horizontal boreholes.

Jan 1, 1983