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The Iron and Steel Committee of the American Institute, of mining and Metallurgical Engineers held a joint session with the American Foundrymen's Association during the Annual Meeting of the Institute in February, 1928, for the purpose of discussing blast-furnace operations and the character of pig iron and castings. Dr. Richard Moldenke presided. H.. H. Sweetser, who was named on the program as the presiding officer of the session, was unable to be present, but sent a letter in which he said: " The meeting we had here last year and the meeting held in Chicago last June showed a desire on the part of blast-furnace men and users of pig iron to get the facts regarding the influence on castings made from pig iron produced in blast furnaces under varying conditions of temperature, mixture and local furnace operations. It has been my hope that at this Round Table we might he able to formulate a series of tests that, would give us the desired information. "Several independent blast-furnace investigations have. been, and still are being, carried on in different sections of the country. The furnaces investigated include a southern furnace, a middle-west furnace, a western furnace and a Chicago district furnace. The results so far show distinct characteristics for each of the four furnaces. It is possible that a furnace in the eastern part of the United States would be different from all the others. "It is hoped that a plan of action can be worked out at your meeting today whereby the IT. S. Bureau of Mines will be able, to carry on more extensive work in connection with the research departments of the different iron and steel companies and the big consumers of pig iron. "No one group can get all of the, information but if the different investigations can be coordinated it will he possible for us to (Iran. conclusions regarding the effects of varying conditions in blast-furnace practice on the pig iron produced and the castings made there from." DISCUSSION R. Moldenke, Watehung, N. J.—I was across the -water this summer and saw some of those interested in an investigation of this kind. 'I'hey told me that in Germany, where they had a joint committee of blast-furnace operators and foundry-

By D. J. Blickwede

WHEN iron is dipped in molten zinc at temperatures in the range 840" to 930°F, the two metals react to form layers of Fe-Zn compounds which adhere to the iron. At the galvanizing tem- perature, the successive layers are:' y (Fe5Zn21), 61 (FeZn;), and 5 (FeZn13), in conformity with the Fe-Zn constitution diagram. There is no quantitative information in the literature on the rates of growth of the individual alloy layers over the temperature range 840° to 930°F. Rather all the past experiments have studied the loss in weight of iron as a function of galvanizing

Jan 1, 1954

By Mason J. Christner

Geology has become increasingly quantitative in the last 20 years. This trend has been accelerated by computer capability in the manipulation of data. The mining industry has been relatively slow to harness computer power, but the larger companies have moved in this direction in the past few years. Among the leaders in this progression is the White Pine Copper Co., and the following is a brief description of their computer system for processing, managing, and utilizing geologic data. Geology The White Pine mine is in the Upper Peninsula of Michigan, 6 miles south of Lake Superior and 80 miles west of the Houghton-Hancock area, center of the famous Copper country. The ore body of chalcocite, native copper, and minor native silver, is a tabular body in the basal 25 ft of the flat lying Nonsuch Shale and the top part of the underlying Copper Harbor Conglomerate. Twelve or more

Jan 1, 1972

By Julian M. Avery

In its dual role of pig-iron smelter and gas producer, the blast furnace is a remarkably satisfactory and efficient apparatus. Many metallurgists and engineers have pointed out, however, that since the primary function of the blast furnace is to produce pig iron at the lowest possible cost, its function as a gas producer should be subordinated to this primary purpose. During the past 10 years or more, improvements in blast-furnace technique and economy have, in fact, been directed primarily to increasing efficiency in the use of the reducing power of the gases produced by combustion of coke at the tuyeres, to reduce iron ore within the furnace. The net over-all result is a decrease in the fuel value of the top gas per unit of volume, and per ton of pig iron, which is evidently more than offset in value by the corollary saving in coke and the possibility of increased furnace capacity, for otherwise there would be no point in making such improvements. The purpose of the present paper is to suggest that, while progress has been made in this direction, there is another approach to the problem of increasing the chemical and thermal efficiency of the blast furnace in its primary role as a pig-iron smelter, which appears to offer possibilities of improvement in over-all ecorlomy of a different order of magnitude than results thus far obtained by the use of other expedients. The present argument is based in large part upon the hypothesis that the efficiency and economics of pig-iron smelting are seriously and adversely affected by the phenomenon known as "solution loss," and that the problem of solution loss call be successfully attacked by increasing the over-all pressure; i.e., the "top prcssure" under which the furnace is operated. It is indicated that pressure operation may also be expected to result in concomitant benefits, which may be summed up as generally improved operating conditions, together with a substantial saving in fuel and/or a substantial increase in furnace capacity. In a broad sense, the suggestion of pressure operation of the pig-iron blast furnace is not new, for operation under moderately increased over-all pressure was suggested as early as 1871, by no less an authority than

Jan 1, 1938

By Julian M. Avery

In its dual role of pig-iron smelter and gas producer, the blast furnace is a remarkably satisfactory and efficient apparatus. Many metallurgists and engineers have pointed out, however, that since the primary function of the blast furnace is to produce pig iron at the lowest possible cost, its function as a gas producer should be subordinated to this primary purpose. During the past 10 years or more, improvements in blast-furnace technique and economy have, in fact, been directed primarily to increasing efficiency in the use of the reducing power of the gases produced by combustion of coke at the tuyeres, to reduce iron ore within the furnace. The net over-all result is a decrease in the fuel value of the top gas per unit of volume, and per ton of pig iron, which is evidently more than offset in value by the corollary saving in coke and the possibility of increased furnace capacity, for otherwise there would be no point in making such improvements. The purpose of the present paper is to suggest that, while progress has been made in this direction, there is another approach to the problem of increasing the chemical and thermal efficiency of the blast furnace in its primary role as a pig-iron smelter, which appears to offer possibilities of improvement in over-all ecorlomy of a different order of magnitude than results thus far obtained by the use of other expedients. The present argument is based in large part upon the hypothesis that the efficiency and economics of pig-iron smelting are seriously and adversely affected by the phenomenon known as "solution loss," and that the problem of solution loss call be successfully attacked by increasing the over-all pressure; i.e., the "top prcssure" under which the furnace is operated. It is indicated that pressure operation may also be expected to result in concomitant benefits, which may be summed up as generally improved operating conditions, together with a substantial saving in fuel and/or a substantial increase in furnace capacity. In a broad sense, the suggestion of pressure operation of the pig-iron blast furnace is not new, for operation under moderately increased over-all pressure was suggested as early as 1871, by no less an authority than

Jan 1, 1938

By Russell G. Wayland

THIS paper gives the author's personal idea of the general viewpoint of the world's largest mica consumer, the U. S. Army, toward the supply, uses, and conservation of mica. However, to cover so broad a subject in a brief paper will require the dismissal of many phases with little or no mention. For example, it will be necessary to omit further reference to phlogopite or amber mica, even though amber block and splittings for certain spark plugs and commutator segments present some supply problems. This paper will deal entirely with muscovite, or white mica. MUSCOVITE MICA Muscovite mica is available commercially in four general forms; namely, as block mica, as splittings, as washer punch, and as scrap mica.1 No strategic importance is attached to washer punch and scrap mica, even though scrap has found some use in military construction and in electrical ceramics. The use of washer punch as such has fallen off with the non-manufacture of household electrical appliances. Some qualities of block mica are also of little or no interest, but other qualities of block mica and all splittings are considered of vital importance. Army equipment requires large amounts of muscovite mica splittings in the form of built-up mica commutator segments and coil insulation in motors and generators. Other military items using significant amounts of muscovite splittings are transformers, switchboards, and blasting apparatus. Flexible mica tape, made from bookform or loose-dusted splittings cemented to paper, linen, glass cloth, and other backing, is widely used in coil insulation in heavy electrical equipment and in aircraft generators. All muscovite mica splittings come from India, principally because of the enormous amount of hand labor required in their preparation. In this emergency, the need for them was anticipated and large industry and government stocks were on hand, while shipments have continued to come by sea in spite of wartime hazards. With the possible exception of certain bookform splittings, no trouble is anticipated in supply. Accordingly, the remainder of this paper is devoted to block (or sheet) muscovite mica, including condenser film that is split from block. THE MICA PROBLEM Block muscovite mica is commercially classified into the following general qualities, beginning with the best quality: I. Clear and Slightly Stained (C & SS). 2. Fair Stained (FS). 3. Good Stained (GS). 4. Stained (S). 5. Heavy Stained (HS). 6. Black Spotted. The ratio of relative commercial abundance in the present new supply of these micas is approximately I C & SS to 4 FS to 10 GS to 14 S to whatever the

Jan 1, 1944

By S. H. Ash

SAFETY records have improved in all branches of the mineral industry. While annual production was rising from $2 billion in 1910 to nearly $12 billion in 1950, fatalities decreased from 3539 in 1911 to 974 in 1951. And, although safety progressed under hand methods, it progresses more effectively under modern, mechanized methods. But powered devices such as electric or diesel units can in themselves create new hazards unless proper safeguards are instituted. Health Use of diesels underground is increasing. Diesel equipment has been used in construction of 34 tunnels in 16 states in the last 17 years. The 77 noncoal underground mines in 24 states using diesels underground range from mines with large open stopes to those with narrow veins, so that the ventilating problems represent a good cross-section of U. S. mining. The U. S. Bureau of Mines has received no report that anyone has suffered ill effects from the proper use of properly maintained equipment. The necessity for correct maintenance and use of satisfactory exhaust conditioners cannot be emphasized too strongly.

Jan 2, 1954

By P. C. Johnson, B. A. Stein

This paper describes a study of the effects of combined heat treatment and explosive loading on the mechanical properties of high-strength steels. nis program investigated two distinct areas: 1) the effect of shock waves, without gross irreversible defmmution, on a 3-Cr steel at various stages of heat treatment; and 2) the effect of rapid deformation (explosive forming) on H-11 and D6-AC steels in the metastable austenitic state. The mechanical properties of these steels were improved, in some cases markedly, as a result of these treatments. ,AUSFORMWG, which requires the plastic deformation of metastable austenite, is a process which can appreciably improve the properties of selected alloy stee1s.l,2 The Ausform process significantly increases the strength of these steels without decreasing their ductility. The properties at high temperatures are also improved through a change in the response of the steels to tempering. Although the mechanism by which ausforming alters the properties of these steels is not fully understood, it appears that the dislocation arrays produced by deformation of the metastable austenite influence the structure of the martensite on subsequent transformation. This, in turn, affects the strength, ductility, and tempering response of the martensite. This research used chemical explosives to deform steels at various stages in their heat treatment in order to improve the properties of these steels. The explosive energy is used in two ways; 1) high-pressure shock waves are propagated through the steel to produce extensive microscopic shear strain without causing a large irreversible change in shape, and 2) explosive energy is used to cause extensive macroscopic plastic strain in the metastable austenitic state (explosive forming). I) AUSFORMING WITH INTENSE SHOCK WAVES The steel used in this phase of the research was an alloy having a nominal composition 0.43 pct C, 3.0 pct Cr, 1.5 pct Ni, and 1.5 pct Si. The steel was subjected to intense shock waves in three conditions: 1) in the metastable austenitic state, 2) in the tempered mar tens itic state, and 3) in the tempered martensitic state after ausforming by conventional techniques. The specimens were in the form of disks 2.75 in. in diam and 5/16 in. thick. These were incorporated into a specimen assembly consisting of two disks pressed into a 5 by 5 by 1 in. block of stainless steel, Fig. 1. Spalling (or scabbing) is confined to the front disk. The specimen is protected from oxidation and decarburization by the surrounding metal. The temperature of the assembly is monitored by a thermocouple inserted into one side of the stainless steel block. The assembly is positioned over an oil reservoir which serves both as a means of catching the disks and as quenching medium for the disks shocked under ausforming conditions. Plane shock waves are introduced into the assembly by a metal driver plate impacting the top surface of the block. The driver plate is accelerated by a chemical explosive sheet supplied by E. I. du Pont de Nemours & Co. All the specimens were subjected to plane shock waves having a peak pressure of approximately 430 kbar. The pressure is that quoted by G. E. Dieter for the plane wave generator used in this work.' The driver plate used was 1/4 in. thick, so that the initial pulse was essentially a l/2-in.-wide square wave. The attenuation of the peak pressure during the subsequent 1/4 in. is estimated to be less than 5 pct. The shock front induces a temperature rise, a portion of which is irreversible. Rough estimates (+25 pct) of this temperature rise have been made for iron shocked at room temperature.4 For a 500-kbar shock wave, the temperature rise in the shock front is about 700°F, and is held for a time of the order of microseconds. The irreversible temperature rise, which remains after the shock wave passes, is about 450°F.4 The disks are quenched to room temperature within a few seconds of the shock treatment. It should be emphasized that the temperature rises given above are estimates for pure iron at room temperature, and are not necessarily true for the tests made in this work. The disks shocked at temperatures in the metastable austenitic range were austenitized in the stainless steel assembly in a furnace protected from the firing area. The assembly was removed from the furnace and placed over the recovery reservoir. The plane wave generator was then positioned and

Jan 1, 1963

By G. F. Huff, G. R. Bailey, J. H. Richards

An improved bomb-sampling technique for obtaining samples for oxygen analysis from liquid steel is described. Analyses of samples taken from open-hearth furnaces by the improved method show sufficient agreement with laboratory data to indicate the accuracy of the method is better than that of other existing sampling methods. THE bomb-sampling method of determining the oxygen content of a liquid-steel bath was studied by obtaining and analyzing more than 200 samples from normal open-hearth heats. Of these samples, about 60 were obtained with a bomb of conventional design;' the remainder were taken with a somewhat modified bomb designed to give greater protection against slag contamination. In the new device, the contacting faces of the cast-iron mold halves are machined so that they fit closely; the opening at the top of the bomb is covered with a thin steel cap; and the top of the bomb is further covered with a wood block held in place by heavy steel wire. After the bomb has passed through the slag layer and has entered the steel bath, the wire melts and the wood block floats to the surface. Thereafter, the steel melts through the thin cap and flows into the bomb cavity where it is deoxidized with aluminum wire. An assembled bomb, ready for sampling, is shown in Fig. 1. To standardize sampling procedure, the samples were usually taken through the center door and at the bottom of the furnace; that is, the bomb was on the bottom when the metal cap melted. After each addition to the furnace, at least 20 min elapsed before a sample was taken. The oxygen contents of all the samples were determined by using the vacuum-fusion method, two or more specimens from each sample being analyzed. The specimens were cubes of metal cut from a region in the interior of the bomb sample, as shown in Fig. 2. Drillings for carbon analysis were obtained immediately above the oxygen-analysis region. Reproducibility of Results In determining the reproducibility of the results obtained with the modified bomb, the following four factors were studied: 1—variation in the oxygen concentration of duplicate specimens taken from a single sample; 2—variation in the oxygen concentration of duplicate samples taken in quick succession: 3—effect of the amount of aluminum wire

Jan 1, 1953

By Nicholas J. Grant, W. Michael Yim

The creep-rupture properties of nickel, in as-prestrained or prestrain-polygonized condition, were studied at 1300°F and 4000 psi, and also at 700°F and 26,000 psi. An improvement of strength was noted in both the as -prestrained and the prestrain-polygonized tests at each temperature. The as-prestrained tests, however, showed superior creep resistance to the prestrain-polygonized ones. From the results of X-ray and etch-pit studies, the strengthc gain was interpreted in terms of dislocation density and the stability of preformed substructures. COLD work is known to improve the strength properties of metals at low temperatures. At high temperatures, however, the strength improvement is often lost if gross structural changes, such as re-crystallization, take place prior to or during the test. Even in the absence of recrystallization, a prestrained metal may undergo structural changes— polygonization and the formation of well-defined subgrains—which importantly effect the resultant properties. The relationship between substructure and creep properties has received considerable attention in recent years. The work of wood1 indicated that the creep strength of aluminum and its alloys was closely related to the subgrain size. Hazlett and Hansen,2 and Ancker, Hazlett, and parker,3 working with nickel and nickel alloys, had shown that small amounts of prestrain at room temperature followed by a recovery anneal at 800°C (1470°F) improved the strength at both low and high temperatures. The strength gain was attributed to an increase in the number of subgrain boundaries as a result of the prestrain-polygonization treatment. More recently Coldren and Freeman4 studied the effect of substructures on the creep rate of nickel at 1550°F. Their conclusions indicated that the substructure size as measured by the number of boundaries was the factor that controlled the creep rate. In addition, the perfection of the subgrain boundary has been indicated to have an important role on the strength.' On the other hand, there are data which disagree with the above observations. Grant, Chaudhuri, Silver, and Ganow6 found no creep strength improvement by a prestraining treatment. In reviewing the role of subgrains on high-temperature creep, Shepard and Dorn7 concluded that the contribution of the subgrains to high-temperature creep strength was a minor one. Guard,8 in a recent study on poly-gonized nickel, has cast doubt on the validity of the postulate that subgrain size controls the creep rate. In the majority of the previous studies, the subgrain boundaries have been singled out as the most important factor in determining creep strength; however, there is no firm theoretical ground for the postulate that the subgrain boundaries act as effective barriers to dislocation motion.B910 Little effort has been directed toward the study of the relationship between creep strength and other structural variables, such as dislocation density, lattice strain,

Jan 1, 1963

By W. J. Schuster

Safety in coal mines depends largely upon adequate training of the foreman. Although management must provide modern and safe equipment and at all times keep mines in first class condition from a safety viewpoint, final results will be determined by the quality of supervision. HANNA COAL CO., Division of Pittsburgh Consolidation Coal Co., operates three large underground mines in eastern Ohio. The section of Pittsburgh No. 8 coal seam in which these mines are located varies in thickness from 52 to 64 in. It is immediately overlain by a stratum of shaly material 12 to 15 in. thick locally known as draw slate, which is structurally very weak and which disintegrates rapidly upon exposure to atmosphere. Immediately above the draw slate as it is normally found is a band of extremely high ash material 6 to 12 in. thick known as roof coal or rooster coal, and above this is a stratum of conglomerate material varying from 4 to 10 ft in depth. Overlying the conglomerate is a relatively thick stratum of limestone, the first stable material above the Pittsburgh coal seam in eastern Ohio. With the method of full-seam mining that has been adopted, draw slate is shot down, loaded with the coal, and removed in the preparation plants. The roof coal then becomes the permanent roof. The major problem in mining the No. 8 seam in eastern Ohio is control of the roof. Since the strata above the draw slate contains no material with a structure firm enough to provide self-support, the roof begins to sag in a relatively short time after the coal and draw slate have been removed. The problem thus becomes one of getting temporary safety posts under this roof as quickly as possible to prevent a break or separation from occurring either in the roof coal or in the conglomerate above it. Haulage System The Pittsburgh No. 8 seam in eastern Ohio is relatively level, with only minor local dips. Throughout the Hanna Coal Co. mines, entries are generally 12 ft wide. Rooms are driven on a 60" angle on 30-ft centers and are 22 ft wide. No attempt is made to extract the 8-ft pillars between. The entire length of main line haulage is gunited in one mine, and a major portion in another. Two of the mines have single-track main haulage roads with passways. The third, a new mine, is double-tracked, and the roof is supported by steel crossbars, 60 lb or heavier, spaced on 4-ft centers and lagged. In recent years timbering on main line and secondary haulage roads has been accomplished by one of two methods: 1—crossbars are supported on a small section of post set in a hitch hole in the rib, or 2—or a hole is drilled in the rib about 12 in. below the roof, of sufficient depth to fasten securely a short length of 40-lb rail, the bottom of the rail facing the roof, on which a short post is set directly under the crossbar. At present the hitch-hole timbering method is favored. At two of the mines the main line haulage locomotives are 26-ton, 8-wheel units. These locomotives are of the axleless type, each wheel being individually mounted on the frame. The motorman's compartment is encircled by 3-in. armor plate for the protection of the occupants. At the third underground mine conventional 15-ton locomotives are being used. However, these locomotives have been completely rebuilt in the company's shops. Equipment has been streamlined and quarters have been provided for two people, who are protected by heavy steel plate in much the same way described above. This modernization program has been completed on all secondary haulage locomotives at the three mines, and the company is well on the way to similar equipment of the 6-ton section locomotives. The following additional features have been included in their modernization: 1—additional support for the motors to prevent their falling to the middle of the track and derailing the locomotives should a break occur in the suspension bar support; 2—installation of additional bracing to prevent brake rigging from becoming displaced and causing derailments; 3—enclosure of all electric wiring in conduit or raceway; 4—provision of an enclosed compartment for the storage of re-railers, jacks, and other equipment, so that they need not be carried on the outside of the motor; and 5—redesign of the end of the locomotive opposite the operator's compartment to prevent anyone's mounting from that direction. It is interesting to note that some

Jan 1, 1955

By Edward N. Trump

EXTENSIVE beds of rock salt occur in New York, Michigan, Kansas, and Texas. Wells are drilled through the beds, cased, and equipped with a suspended center tube. By circulating water through such a well, the soluble saline is dissolved and a brine produced. Operation of a well results ordinarily in the formation of an inverted conical cavity in the saline, inasmuch as the dissolving action of the fresh water is more rapid at the top of the bed than at the bottom. Under continued operations, the side wall of the cavity becomes so flat that a blanket of mud is deposited thereon and if the bed is overlain with shale or a soft rock caving follows. Three methods of operating wells have been employed-the "Tully plan," "the Detroit plan," and "the Trump plan." Both the Tully plan and the Detroit plan involve the formation of the aforesaid conical cavity with constant caving of shale above and deposition of mud on the side wall; the Trump plan was developed to avoid both these circumstances and thus materially prolong the life of a well. Details of the three plans follow. THE TULLY PLAN In the Tully plan, the water passes down the annular space between the casing and the center -tube, the brine rising in the center tube (Figs. I and 2). An inverted conical cavity is formed, the side wall of which relatively soon becomes so flat that it is blanketed by mud and impurities from the rock salt; this blanket limits dissolving action to the perimeter of the base of the cone, situated at the top of the saline bed. Caving of the overlying strata follows, and the total production of brine from the well is comparatively small. THE DETROIT PLAN In the Detroit plan, the water passes down the center tube, the brine rising in the annular space between the center tube and casing (Figs. I and 2). Initially, this flow forms a pear-shaped cavity in the bottom of the salt bed, but this cavity soon changes in shape to that of an inverted cone, as under the Tully plan, with the attendant disadvantages inherent in that plan. Application of the Detroit plan is limited to salt beds 300 ft. thick as a minimum. This depth of bed is necessary to prevent fresh water of lower specific gravity than the brine from rising in the

Jan 1, 1944

By C. T. Holland

This discussion is concerned with those com-J- paratively infrequent bumps that eject material from the failed mass with enough energy to wreck heavy machinery and seriously injure or kill people. In such cases there is a loud report, and the pressure waves set up in the mine atmosphere wreck ventilation controls several hundred feet from the site of the failure. Vibrations are induced in the earth's crust which may be detected several miles from the point of origin, and explosive gases may be released in large quantities." Usually a dense suspension of dust is produced from the failed mass, together with dust placed in suspension by the resulting air blast. These rock bursts or bumps have also been called mountain shots, bounces, pillar bursts, and crumps. Such occurrences, when extensive, are serious catastrophes, and may force abandonment of the mine.' CAUSES OF ROCK BURSTS Rock bursts in coal mines have occurred in sev-eral places under diverse geological conditions, mining methods, and/or practices. A study of many such occurrences indicates that they depend on the following conditions: 1) cover thickness of 500 ft or more, 2) cover composed of strong members and having a strong member lying close to or on the coal bed, and 3) a floor that does not readily heave. The minimum thickness of cover under which coal mine rock bursts occur is based on observation. Under unusual circumstances involving very strong roof members and/or mining systems, and/or practices especially likely to induce bursts, a burst conceivably could occur under less roof. Under special circumstances rock bursts have occurred in marble quarries almost at the surface.' Furthermore, under very heavy cover it is likely that controlling roof members need not be as strong and as close to the coal bed to cause bumps. Just what constitutes a strong roof or roof member is open to question. Obviously the strength of a roof member is determined by its thickness, the distance between bedding planes across which little or no bond exists, the spacing and development of joints, and the inherent strength of the rock material. Strata sections showing roofs under which rock bursts in coal mines have developed are illustrated in Fig. 2. In several instances the strength of rocks composing the roof of mines developing bumps has been tested. The tests (Table I) show that these rocks develop considerable strength in compression, bending, and shear and that in some cases the strengths are very high. No quantitative data are available concerning joints in roof members over coal beds where rock bursts have occurred, but the author has examined many mines subject to bursts and has received the impression that joints in the shales immediately over the coal beds are always present and well developed. In the main sandstone or conglomerate members themselves, however, where observed, he has yet to see a fracture system indicating well developed joints. Undoubtedly joints are present, but the evidence indicates that they are few, poorly developed, and widely spaced. In a mine subject to bumps, moreover, actual measurements of roof spans up to 225 ft between supports under a cover of about 800 ft have been made before roof failure occurred. Such spans signify strong roof. In coal beds subject to rock bursts, the bottom rock is usually a sandy shale. A few tests (Table I) have been made on the floor rock of coal beds. Observations in mines indicate that they are usually weaker than roof rocks and that under load they show more plastic effects. Observation in ground subject to bursts invariably indicates a relatively strong floor highly resistant to heaving, although some heaving may occur. Other factors listed as favorable to rock bursts are: 1) mountainous surface; 2) steeply dipping beds; 3) proximity of faults, folds, and other similar geologic structures. Many bursts have occurred in mines under mountainous country, although many bursts in mines under a surface with little relief indicate that thickness of cover, not relief, is the essential factor. In steeply dipping beds, many bursts occur in mines that are essentially flat, and rock bursts have been recorded in beds dipping as much as 35o Most rock bursts in the U. S. occur in beds that dip less than 10 pct. It is also true that most U. S. coal is produced from beds dipping less than 10 pct. Geologically, faults are regarded as providing stress relief, but in the vicinity of faults, as well as sharp folds, it would seem that residual stresses could be present. The fact that rock bursts occur in areas remote from faults and sharp folds suggests that their presence is not necessary for this phenomenon to take place. There have been many bursts close to faults and folds, and it may be assumed that occasionally such structures accentuate their occurrence3 The conditions and circumstances under which they occur, coupled with the effects produced, leave In the Crowsnest Pass coalfield, British Columbia, a rock burst is estimated to have released 1.8 million cu ft of methane.

Jan 1, 1959

By F. M. Morris

A large reserve of thick coal in southwest Virginia was developed by Clinch-field Coal Co. in 1957-1958 to produce a nominal rate of 1500 tph raw coal. Operation features coal cleaning in transit. Refuse removed averages 35 pct. Evolution of plant from initial conception to completion is discussed, selection of means applied is explained, and performance to date us design expectation is described. The long-range plan calls for ultimate handling capacity of 40,000 tpd raw coal with anticipated clean coal capacity of 26,000 tpd on a three-shift basis. From a materials handling viewpoint, the Moss No. 3 operation is principally of interest as an ensemble. Generally speaking, it uses time-tested equipment and ideas but some of these are employed on a scale that may be new in the industry. At present about 20,000 tpd of raw coal are being handled. This is expected to increase to 40,000 tpd as soon as business conditions justify it. The purpose of this paper is to discuss the evolution of the plant as to materials handling practices, to describe briefly the equipment and methods used, and to comment on performance in certain areas. The subject divides itself naturally into four phases: 1) operations at the mines, 2) transportation from mines to plant, 3) raw coal handling into the plant, and 4) transportation of refuse away from the plant. OPERATIONS AT THE MINES The coal reserves for Moss 3 are in southwest Virginia where Dickenson, Russell, and Buchanan Counties come together. This area contains about 15 square miles and over 100,000,000 tons of coal. Here the No. 4 (Tiller) and No. 5 (Jawbone) seams of the Norton Formation lie so closely together that for practical purposes, they constitute one seam of coal. This seam, which is called the Thick Tiller, varies from 10 to 18 ft in thickness and underlies Sandy Ridge, a mountain cresting between 2400 and 3300 ft above sea-level. At 18-ft seam height (which is considered to be the maximum practical mining height) the parting between seams will be about 3 1/2 ft thick and each bench of coal will be about 7 ft thick. Depending on the amount of impurities in the seams Drover. total reject in this height coal may approximate 50 pet by wt. It will average about 35 pet for the property as a whole. The first move in developing this resource was made in 1953 when contour maps of several square miles around Duty, Va., including all the known outcrop, were made by photogrammetry. At this time, it was felt the prospective operation would be served by the Clinchfield Railroad and a photogrammetric route survey was made by this railroad from Haysi to Duty. Study of the resulting maps indicated only one site—adversely owned—which might accommodate the size washing plant to be erected. Water resources of a dependable nature seemed nonexistent. In 1954 bulk washability tests were made on the Tiller Bench at the Moss No. 1 preparation plant. The tests indicated this portion of the seam, mined separately, would wash to 4 pet ash with good recovery. Also in 1954, development of Moss No. 2, south of Sandy Ridge, was begun. This mine is in the Tiller Seam where it is about 100 ft below the Jawbone Seam. The reasons for developing a mine in the normal Tiller Seam before tackling the Thick Tiller seemed compelling: Railroad service could be established quickly, communications were better (though not good!), more was known of the seam (it had been mined in the years 1911 to 19241, and there was no essential property to be acquired. After some legal skirmishing, the Norfolk & Western Railroad was granted the right to serve the new mine. Three decades earlier the old mine had been served by the Clinchfield Railroad. The event which triggered active development of Moss 3 was the Appalachian Power Co.'s decision in 1956 to build a 450,000-kva power plant on Clinch River at Carbo. This solved the problem of marketing the steam coal which inevitably must be a product of a mine in the Thick Tiller. Management promptly decided to build the preparation plant at Carbo where an excellent site was owned; where railroad service existed; where telephone service could be obtained; and where roads, bridges, and water supply were tolerable.

Jan 1, 1961

By Jean-Jacques Prompsy, J. S. Rinehart

The purpose of the present study was 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 drill 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. This paper covers one phase of a broad study 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, arange commonly encountered in practical drilling operations; and impulsive loading through the detonation of explosives, a velocity range in which the dynamic effects are greatly exaggerated and made more identifiable. Specifically, fracturing which develops at the end of a cylindrical bar undergoing a shock produced by impact was experimentally investigated, the impacted area being the face of a truncated cone terminating the cylindrical bar. Fracture characteristics were looked at as a function of impact velocity, area of impacted face and cone angle — parameters of considerable importance in the design of drill bit. EXPERIMENTS The specimens were made of Plexiglas, which enabled the fracture pattern to be easily observed. In Plexiglas, the velocity of a longitudinal wave is about 8,700 ft/sec, and the velocity of the transverse wave about 4,200 ft/sec. The specimens were impacted by a steel cylinder whose lower face was parallel to the terminal face of the experimental bit, as shown in Fig. 1. The impacting mass was shifted up and down on two steel rods, dropping from a maximum height of 11 ft, with the velocity varied between 0 and 6.5 m/sec. The initial mass of the cylinder was 4.1 kg, which could be increased by steps up to 8.8 kg by addition of steel plates. The velocity of the steel cylinder was determined by using an electric circuit; as the cylinder dropped, it contacted two steel wires, the first one starting and the second one stopping an electronic counter. Two similar counters connected in parallel were used to corroborate each other. The velocity was checked at each drop as it had been observed to vary somewhat from drop to drop. The two contact wires were located just above the top of the specimen so as to give a value of the instantaneous velocity of the steel cylinder at impact. The distance between the wires was 40 ± 0.5 mm and the discrepancy between the counters was never more than 0.1 m/sec. Thus the velocity was known to within 3.25 per cent. This precision was sufficient with respect to the precision of the machining of the specimens. A very small lack of parallelism between the face of a specimen and that of the impacting hammer resulted in an oblique impact which required a much lower velocity to break the specimen. The experiments had a twofold purpose — (1) to observe the fracture pattern, and (2) to find relationships between the minimum velocity required to generate a fracture within a specimen and the

By C. B. Sclar and

A graphical solution is presented for the equation v =v &-cU where vb is the volume of liquid b of density 6 that must be added to liquid a of volume va and density d, in order to obtain a heavy-liquid solution of preferred density dm for mineral fractionation. The equation is valid only for pairs of liquids whose volumes are additive, but empirical density-composi-tion data show that this condition is met over wide compositional ranges by all of the heavy solutions that are commonly used for mineralogical analysis. The chart is a nomograph which consists of two horizontal volume scales ad a vertical density scale arranged so that one volume scale occurs on each side of the density scale. All the scales are linear. For two liquids of density d, and &, respectively, each volume scale has an independent level on the density scale, ad the graphical solution for vb on the nomograph is obtained by construction of one straight line. The chart can be prepared easily to cover any density range with any desired accuracy by proper selection of the scales. Heavy-liquid separation of particulate samples is an indispensable analytical procedure in the mineralogical and mineral-process ing laboratory. Heavy liquids may be used 1) to isolate specific minerals in the form of high purity products or to fractionate samples into several products with density limits for petrographic, chemical, spectrographic, and X-ray diffraction analysis,'-' 2) to determine the density of minerals,'-l4 3) to facilitate the recognition and identification of optically similar associated minerals,15 and 4) to determine quantitatively the degree of liberation of ore minerals in ores and mill products at various limiting sizes." The preparation and recovery of heavy liquids suitable for mineralogical analysis are discussed fully in the literature and operational procedures and special equipment for heavy-liquid separation are described in numerous references.'747837' 13,18,25-42 Useful tabulations of heavy liquids for mineralogical analysis are given by Tickell, Mil-ner,' Twenhofel and Tyler,' and Lange and a very complete up-to-date list of minerals arranged according to increasing density is given by Mursky and Thompson. USEFULNESS OF THE DENSITY CHART For any particulate sample, the limiting densities selected for its fractionation by means of heavy- liquid separation depend on the respective densities of the constituent minerals and the specific objectives of the separation. In many instances, the densities of readily available and practicable heavy liquids do not coincide with those that are required, and suitable liquids of the proper density are obtained by dilution with miscible liquids. In the general case, one may wish to mix either a pure liquid (A) or a solution of two miscible liquids (AP,,) of known density with either a pure liquid (B) or a different solution of the two miscible liquids (A,B,) of higher or lower density in order to obtain a final liquid of the preferred density. In conventional practice, the latter is reached empirically by trial-and-error addition of increments of one liquid to the other. The density of the final solution may be determined either by accurately weighing a known volume of the final liquid,'' 532;4;40, or by means of I)a Westphal balance, 2)a hydrometer, a refractometer in the case of binary liquid systems for which density-refractive index data are available,15 4) natural or synthetic .indicators of known density, a pycno-meter. Like all trial-and-error methods, these procedures are time consuming and many require special equipment which may not be available, The purpose of this paper is to show that the volume of a liquid of known density which is to be added to another liquid of known volume and density in order to obtain a solution with an intermediate preferred density may be determined rapidly and accurately by a graphical method provided that the two liquids form ideal or quasi-ideal solutions whose volumes are additive. Empirical data47'8~'8~4B'49 show that solutions prepared from pairs of miscible

Jan 1, 1961

[Year of Election to Honorary Membership 1876. PROF. RICHARD AKERMAN Stockholm, Sweden. 1913. DR. FRANK DAWSON ADAMS Montreal, Canada. 1920. HENRY STURGIS DRINKER Bethlehem, Pa. 1888. PROF. HATON DE LA GOUPILLIERE Paris, France. 1906. SIR ROBERT A. HADFIELD London, England. 1917. HERBERT HOOVER Stanford University, Cal. 1919. ROBERT W. HUNT Chicago, 111. 1915. JAMES FURMAN KEMP New York, N. Y. 1905. PROF. HENRI Louis LE CHATELIER Paris, France. 1913. EZEQUIEL ORDOENEZ Mexico City, Mexico. 1909. ALEXANDRE POURCEL Paris, France. 1911. ROBERT H. RICHARDS Boston, Mass. 1906. JOHN E. STEAD Middlesbrough, England. 1902. PROF. DIMITRY CONSTANTIN TSCHERNOFF Petrograd, Russia. 1907. CHARLES D. WALCOTT Washington, D. C.] [ ]

Jan 1, 1925

By Daniel R. Shawe

Ore deposits of the western U.S. are divisible broadly into two genetic groups, a smaller one that fits the plate tectonic model of subduction and a larger one that appears unrelated to subduction. Some of the middle Paleozoic and lower-middle Mesozoic ore deposits in the western U.S. are associated with eugeosynclinal rocks and lie a few hundred kilometers west of reconstructed continent margins. These deposits probably formed as a result of subduction beneath oceanic or thin continental crust that was analogous to subduction predicated for the Kuroko massive sulfide deposits of Japan. Some middle Mesozoic porphyry copper deposits may have formed east of the continental margin when subduction-generated magmas were emplaced high in continental crust in a manner analogous to subduction-generated magmas that formed the Andean porphyry copper deposits. Most Cretaceous-Cenozoic ore deposits lie in a broad region extending many hundreds of kilometers east of the continent margin and characterized by conjugate tectonic zones, high terrestial heat flow, and low seismic velocity in the upper mantle. The deposits likely were formed in relation to magmas generated along the conjugate tectonic zones during differential northwestward drift and internal wrenching of the continental plate.

Jan 1, 1978

By R. H. Frazier, C. H. Lorig, F. W. Boulger

THERE are both advantages and disadvantages in using semikilled steels in place of killed steels. One advantage of semikilled steels is they provide a higher ingot-to-product yield. This is especially important to shipbuilders in the time of national emergency. By using semikilled steel for hull plates, a greater number of ships can be built from the same ingot tonnage. Consequently, if semikilled steel plates are adequate for the service, the advantage of the higher product yield is of considerable importance. The performance of hull plates is closely associated with the ductile-to-brittle transition temperature, and this, in turn, is strongly dependent on the composition of the steel. A low transition temperature is desirable because it indicates that the steel is less likely to fail at low ambient temperatures. Killed steels are known to have lower transition temperatures than semikilled steels. It is believed that the better qualities of killed steels in this respect are due mainly to the low oxygen contents of the steel. The principal deoxidizers, aluminum, silicon, and manganese, lower the oxygen content. Fundamental studies1-0 have shown that the oxygen content remaining after the addition of one of these three elements is influenced by the residual amounts of the other two present. In the current study, therefore, various amounts of silicon and aluminum were added to steels containing different manganese contents for the purpose of studying the influence of silicon and aluminum on the notched-bar properties of hot-rolled steels. Eleven types of steels were studied. The temperature at which the plates were finish-rolled was care- fully controlled at 1850°F. All plates were rolled to ¾-in. thickness, and tested in the as-rolled condition. The total number of steels prepared in the laboratory for this study was 95. The charge consisting of ingot iron and ferrosilicon equivalent to 0.10 pct Si was melted in a 200-lb induction furnace under an atmosphere of argon to insure low, uniform nitrogen contents. After the charge was melted and the desired temperature was obtained, the melt was partly deoxidized with either silicomanganese or aluminum. Aluminum was used for this purpose in only those steels with very low silicon contents, where the finished steel was to contain some aluminum. This initial deoxidizing addition was made to obtain consistent recoveries of subsequent additions of ferro-manganese, ferrosilicon, and aluminum. Carbon, in the form of graphite, was added, either just prior to tap or to the final aluminum addition. The entire heat of 200 lb was poured directly into a 6x6-in. big-end-up mold. The ingots of semikilled steel were capped with a steel plate. The killed steels, on the other hand, were poured with a hot-top containing 14 pct of the total volume of the ingot. The ingots were processed by heating to 2250°F, followed by forging to slabs 1¾ in. thick and 6 in. wide. After reheating to 2250°F, the slabs were rolled to 0.9-in. gage, using reductions of approximately 1/6 in. per pass. The 0.9-in. thick plates were immediately recharged in a furnace held at 1850°F. After 30 min in the furnace at 1850°F, the plates were rolled to ¾ in. in one pass. Following the final pass, the plates were stacked on edge on a brick floor, with a brick separating one from another, where they were allowed to cool in air. Drillings for chemical analysis were taken from the plates at locations corresponding to the top and bottom of each ingot. Carbon, manganese, silicon, phosphorus, sulfur, and nitrogen contents of each sample were determined. The analysis of each steel was carefully controlled so that the contents of

Jan 1, 1957

[Year or HONORARY MEMBERS Election 1876. PROF. RICHARD AKERMAN Stockholm, Sweden. 1913. DR. FRANK DAWSON ADAMS Montreal,. Canada. 1888. PROF. HATON DE LA GOUPILLIERE Paris, France. 1906. SIR ROBERT A. HADFIELD London, England. 1896. HERBERT HOOVER Stanford University, Cal. 1874. ROBERT W. HUNT Chicago, Ill. 1915. JAMES FURMAN KEMP New York, N. Y. 1905. PROF. HENRI Louis LE CHATELIER Paris, France. 1918. EZEQUIEL OBDONEZ Mexico City, Mexico. 1909. ALEXANDRE POUBCEL Paris, France. 1911.. PROF. ROBERT H. RICHARDS Boston, Mass. . 1906. JOHN E. STEAD Middlesbrough, England. 1902. PROF. DIMITRY CONSTANTIN TSCHERNOFF Petrograd, Russia. 1910. PROF. TSUNASHIRO WADA Tokyo, Japan. 1907. CHARLES D. WALCOTT Washington, D. C.] [ ]

Jan 1, 1925