Search Documents

By Leon Merker

FLAME fusion growth of strontium titanate crystals was undertaken to obtain large transparent crystals on which physical data could be gathered. The fact that strontium titanate is a cubic crystal and probably has a high index of refraction indicated that it might be useful in optics. No large boules* of strontium titanate had been synthesized previously, and none have been found as natural occurrences. The flame fusion process, invented by Verneuil in 1904,1 was first applied to grow large crystals of ruby and has since been used extensively to grow crystals of corundum and spinel on a commercial scale. Only during the past nine years has this technique been applied to the growth of large single crystals of other high melting compounds such as rutile,2 mullite,3 and ferrites4 One of several significant advantages of the flame fusion technique over more conventional methods of growing crystals from a melt is that the crystals are grown without the use of melt containers. This fact alone excludes other methods of growing crystals from a melt whenever there are compounds with a melting point near 2000°C or above. Growth of the crystal, similar to that of a stalagmite, takes place inside a circular refractory furnace that does not contact the crystal, Fig. 1. This eliminates a serious source of impurities as well as the likelihood of multi-crystalline solidification along the container walls. The flame fusion technique is particularly advantageous for research because it produces large crystals quickly and at lower cost. A corundum crystal 25 mm long and 12 mm in diam can be grown in 2 to 3 hr. Operating conditions permit observation of the growing crystal during any stage of the process. The various flame patterns and a wide range of oxidation-reduction conditions can be investigated in a relatively short time. It should be pointed out that formation of the foot of the boule is the most critical step in growing an unfractured boule. When the tip of the sintered cone is fused to initiate crystal growth, the original area of fusion should be kept as small as possible so that one crystallite can assume dominance over all the others and propagate as a single crystal. If the boule start happens to consist of several crystals, usually the entire boule also contains several crystals. In many cases these tightly cohering crystals exhibit different crystallographic orientations. Although the basic technique of the original Verneuil process has been retained, refinements in method and burner design have been developed in recent years. For example, boule growth may be initiated on a seed crystal rather than by fusion of a sintered cone. The danger of multicrystalline or twinned boule starts is thus eliminated, and this procedure may also be used to grow boules of predetermined crystallographic orientation. Prior Art: Of the many attempts to produce artificial gem crystals, most early efforts% ere directed toward synthesis of .rubies and sapphires and diamonds. None of these efforts resulted in a process that was commercially practical. Without exception only tiny crystals of corundum were obtained, and those at excessive cost. Verneuil announced his process in 1904; Ravier6 in 1934 presented a detailed description of a commercial boule production plant. Descriptions of the German synthetic stone industry, the process details, and the equipment used have been published by Barnes7 and by Merker.'

Jan 1, 1956

By N. A. D. Parlee, I. D. Shah

The solubilities of oxygen in liquid Ag-Au, Ag-Pt, and Ag-Pd alloys have been determined in the range of 940° to 1200°C at 1 atm pressure of oxygen using an improved Sieverts technique. The additions of gold (0 to 57 pct), platinum (0 to 27 pct), and palladium (0 to 22 pct) to liquid silver continuously decrease the solubility of oxygen in liquid silver. Activity coefficients and interaction coefficients of oxygen in these alloys and have been calculated. The interaction coefficientscan be described by the following equations: Approximate second-order interaction coefficients (pM have been calculated. The contribution of the product, pMxNM, to In yo only becomes significant at high alloy concentrations. RECENTLY several investigators1"3 have redeter-mined the solubility of oxygen in liquid silver. In general the recent findings1" for Po, = 1 atm are in good agreement with the equation:4 log %O = 695.4/T + 1.086 [1] It is established1, 2 that the solubility of oxygen follows Sieverts' law reasonably well, at least up to 1 atm partial pressure of oxygen. It should be noted that a very slight deviation from Sieverts' law has been reported by Lupis and Elliott.3 Only recently has the solubility of oxygen in Ag-X binary liquid systems received much attention, and only the Ag-Cu, Ag-Au, Ag-Pt, and Ag-Pd systems have been studied. Par lee, Sacris, and zoellners reported that the solubility of oxygen in liquid silver increases progressively with the small additions of copper (0 to 0.5 pct) up to the point where the copper oxide phase appears. Lupis and ~lliott~ have reported that the solubilities of oxygen in liquid silver decrease progressively with the additions of gold, platinum, and palladium. However, their investigation was limited to 0 to 5 pct of these alloying elements, a range too limited to satisfy the requirements of the research program of this laboratory, where the solubilities of oxygen in the range of 0 to 30 pct were required for the calculation of chemical diffusion coefficients of oxygen in these alloys. The present paper deals only with the solubilities of oxygen in these alloys while another paper will deal with the diffusion of oxygen in these alloys. EXPERIMENTAL METHOD A Sieverts technique was employed which involved an improved design of the reaction tube, Fig. 1, as described briefly in a previous communication.4 Liquid melts (70 to 100 g) were held in Norton Alun-dum crucibles. The induction furnace used had a maximum output of 2.5 kw with very fine control. This induction system had excellent stability and the final equilibrium temperatures measured by calibrated Pt-Pt, 10 pct Rh thermocouples are believed to be within ±2C. These thermocouples were checked against the melting point of pure gold wire, and the results agreed with standard tables (±l°C). Each run involved a careful sequence of operations. The sample was heated to and held at about 900°C in vacuum for 3 to 4 hr with the thermocouple in a retracted position. After melting, the melts were usually exposed to a small amount of oxygen (0.5 to 2.0 cu cm) and then degassed slowly for 1 to 2 min to remove any oxidizable volatile impurities such as carbon and sulfur as CO, CO2, SO2, and SO forth. Argon gas was then allowed to enter the system. The protection tube containing the thermocouple was then

Jan 1, 1969

By E. A. Jones

THE Southeast Missouri division of the St. Joseph Lead Co. normally hauls and hoists over 5 million tons of lead ore each year. This ore is mined in the stopes and headings of 20 mines, hauled to a main line system, see above, and transported up to 6.4 miles to ore-hoisting shafts. Average haulage distance is 2.16 miles. A few years ago a distinguished mining engineer visiting this area for the first time remarked, "You do not have a mining problem; yours is a haulage problem." Its relative importance may be gaged by the fact that mine haulage represents about 25 pct of the total cost of mining ore, the most costly single underground operation. The Southeast Missouri lead belt is a mining district located about 65 miles south of St. Louis. This district is roughly 6 miles wide and 10 miles long. The ore horizon lies about 400 ft below the surface where most of the ore is mined from flat-lying beds of dolomitic limestone. This paper deals with the central part of the district, namely, that portion which contains a group of interconnected mines all in a central haulage system, serving four hoisting shafts. The two major hoisting operations are carried out at Federal No. 17 shaft and Leadwood No. 12 shaft. No. 17 accounted for about 70 pct of the ore hoisted in the district during 1950, averaging about 14,000 tons per operating day. The underground dumping and skip-pocket system at this shaft is similar to those of the other shafts except that it has two rotary dumps. The other ore-hoisting shafts have only one. It might be mentioned that waste rock is hoisted at still other shafts, hauled off, and dumped on the surface by trucks. The hoisting shafts are on a connected system of about 200 miles of underground 24-in. gage main line railroad. There is also the Bonne Terre system, with its own ore-hoisting shaft, which consists of 13.5 miles of 24-in. gage main line, located in the extreme north portion of the lead belt area. Smaller outlying mines such as Doe Run No. 12 and Hayden Creek operate on trackless or shuttle car haulage. In addition to the system of main line haulage, each contributing mine group has its own extensive system of stope or heading railroad. Trackless mining is making considerable headway in the Lead-wood and Desloge mines. Here the shovel loads into a shuttle car which in turn dumps into mine cars spotted on the main line track system. When the merits of underground haulage and surface transportation are compared, it must be remembered that in early days there were many individual

Jan 1, 1954

BETWEEN the third and twentieth days of November, 1903, six intensely interesting letters were mailed from Santiago, Chile, addressed to "Mr. William Braden, Consulting Mining Engineer, 71 Broadway, New York City." The subject of these communications, referred to as the "Rancagua mines," was a copper property situated in the high. Andes; and the writer was Marco Chiapponi, an Italian mining engineer. Many a letter has been written about potential mines and prospects, good and bad, situated in all parts of the world; but few of them combine the shrewdness, sincerity, naïveté, and salesmanship that appear in these fascinating letters by Chiapponi. In 1894, Mr. Braden, as a young man twenty-three years old, had gone to Santiago to look after the erection and operation of some mining machinery that was to be exhibited at the International Mining Exposition. Not the least important of his purposes was to get acquainted at first hand with the mines and mining men of Chile, which ranked at the time (as it does today) second to the United States as a copper-producing country. It happened that a dispute arose regarding charges for certain construction work done in connection with young Braden's exhibit. Finally, arbitration was agreed upon and the Chilean exposition officials proposed Se[n]or Marco Chiapponi as a fair-minded neutral to decide the question. This he did in a manner satisfactory to both sides, and subsequently he and Mr. Braden became good friends. Nine years later Braden wrote to Chiapponi inquiring as to certain matters connected with the Corocoro copper mines in Bolivia. It was to this letter that Chiapponi refers in his letter of November 3, from which the following are some extracts,

Jan 1, 1933

By H. J. Beach

This paper describes the properties and uses of cementitious mixtures containing 10 to 20 per cent bentonite and 2 per cent or more sodium chloride. Salt additions improve bentonite cement by increasing early strength and stabilizing viscosity without changing thickening time. The volume yield of cement is increased 20 per cent or more without lowered early strength. The unique properties of salt in this type of system make savings of more than 10 per cent possible where light-weight, low-strength cementitious material is adequate to fill behind pipe. The density range of these mixtures is 12 to 13 1b/gal and the filtration rates are medium-low, aiding the placement of long columns in primary cementing. Where available, sea water is an ideal salt source in preparing these slurries. INTRODUCTION Accelerators increase the rate of cement hydration. In speeding reaction with water, they can affect cement performance by reducing thickening time and increasing early strength after set. Accelerators that cause both effects can be considered "total", and those which produce only one effect can be called "partial". Materials that act conversely to accelerators are retarders. and they also can be total or partial. Partial retarders. make high-temperature cementing possible by prolonging thickening time while impairing early strength only slightly. Most known accelerators are total. A useful partial accelerator would increase strength without altering thickening time. This report concerns material combinations in which a normally total accelerator exerts unique partial properties. Utilizing these limited effects improves bentonite cement. SEPARATE EFFECTS OF SODIUM CHLORIDE AND BENTONITE ON CEMENT Sodium chloride and calcium chloride are commonly used in cement, and at low concentrations their effects are similar. Ludwig1 studied sodium chloride in API Class A and Class E cements.' Figs. 1 and 2 sum- marize his data taken at 140°F on slurry pumpability and early strength. Normal water-to-cement ratios are 0.46 and 0.40, respectively, for Class A and Class E cements. At saturation the salt content is 12 to 14 per cent of the weight of cement solids. The illustrations, therefore, cover the range of salt solubility in the mixing water. The data show that sodium chloride is a total accelerator of the Class E cement studied over the range of solubility. It is total in Class A cement to 6 per cent. Above this concentration salt becomes a retarder— partial from 6 to 10 per cent and total above. These results are general for different brands of Class A cement, although the shift from acceleration to retardation may occur at different salt concentrations for different temperatures. The results are specific for the Class E brand studied. In discussing his data, the investigator emphasized that different Class E cements might respond erratically to salt, depending on the brand chosen. Morgan and Dumbauld3 escribe the properties of bentonite cement in a paper that is a standard reference in using these systems. Bentonite functions in cement as an enabling agent; each bentonite addition equal to 1 per cent of the weight of dry cement makes possible the further addition of 4.5 per cent water.' Thus, 10 per cent bentonite in Class A cement, normally slurried neat with 46 per cent water, enables the slurry to contain another 45 per cent before settling. Water is one of the lightest and cheapest cement extenders; it reduces density and cost. Bentonite cements produce uniform slurries having low susceptibility to solids separation. After set, they offer low resistance to perforation.

By E. E. Anderson, R. L. Newacheck, J. Kohl

Radioactive materials were used underground before the advent of the nuclear reactor. Radium containing ores have been mixed with cements and other materials for later location in oil wells by gamma ray logging since 1940.1 Perforating bullets have been marked with radium salts,, and radioactive materials have been added to drilling muds to find the cause of "lost circulation."' More recently colloidal suspensions of radioactive cobalt-60 and silver-1103 and solutions of 1-1314 have been used in waterflooding operatiorns to determine injection profiles, flow paths and water velocities. In this paper a recent and interesting underground tracer application is discussed. Development work and field engineering were carried out under the direction of Tracerlab, lnc., with assistance from personnel of the client. A gaseous tracer, argon-41, was used in an attempt to determine the cause of leakage from an underground gas storage reservoir. STATEMENT OF PROBLEM This gaseous tracer application was carried out for the Natural Gas Storage Co. of Illinois. This company has developed an underground gas storage field 1,750 ft below the surface of the earth near the village of Herscher, Ill., about 60 miles south of Chicago. The total volume of gas stored in the field was estimated to be about 13 billion cu ft in the spring of 1954. The field has a maximum estimated storage capacity of 90 billion cu ft. The gas is contained in the Galesville sandstone, previously a water bearing formation of about 25 per cent porosity. The water as it is pushed out by the gas forms a barrier for the gas around the edge of the reservoir. Over the top and beneath the bottom of the reservoir are located several layers of impervious rock, which barring cracks or unplugged and unknown wells should act as seals. Twenty-one injection and withdrawal wells were drilled into the Herscher reservoir and gas injection was begun during the summer of 1953. After a few weeks of gas injection, evidence of gas leakage was detected at the surface; water wells began bubbling and losing prime, and dry spots in the fields indicated gas leakage from uncharted oil wells. In every case, the leakage could not be traced back any further than the water bearing formations a few hundred feet below the surface. An extensive search was made to locate uncharted oil wells which were believed to be possible channels for the gas from the storage formation to the water containing formation. Long-time residents in the area assisted in the location of several abandoned oil wells. The wells when located were cleaned out and either used as vent wells to relieve the gas pressure in the water formations, or were plugged with cement. None of the wells even approached depths which would endanger the soundness of the storage reservoir. All of the injection wells were tested for leaks in the casings by conducting neutron surveys and by other techniques. To detect any leakage up through the cement-filled annulus surrounding the casing, a section of the casing wall in one well was milled out. Then, plugs were installed above and below the cut-out section and pressure applied to the annulus, which proved conclusively that the cement was sound in that well. The well was then plugged with cement and abandoned. In another well, ammonia was injected into the well and a temperature log was made. It was hoped that a leak would be re-

Jan 1, 1956

By W. L. Kerlie, J. H. Richards

The variation in hydrogen content of basic open-hearth steels during refining and the effect of changing from steam to air atomization were studied. The water content of the furnace atmosphere, the slag basicity, and the degree of oxidation of the slag-metal system largely control the hydrogen in the steel. SMALL amounts of hydrogen have an adverse effect on the ductility of steel and a wide range of hydrogen contents has an adverse effect on the internal soundness of both ferrous and nonferrous castings. This has been of major concern to the metal industry, both manufacturers and fabricators, for many years. As one of the results of continuing investigations of the hydrogen problem, much has been learned about methods of reducing the hydrogen content of liquid metals below the relatively high levels that are normally associated with such phenomena as rising ingots and excessive porosity in sand casting. However, it was not until a reliable and simple method for sampling and analyzing liquid metal was developed that much progress was possible in quantitatively evaluating the effect of steelmaking practice on hydrogen contents at levels below 6 to 8 ppm. Throughout all the work described in the following presentation the Geffner pintube technique was used to obtain samples of liquid metal1 * and the tin-fusion method, developed at Massachusetts Institute of Technology,2 was used to analyze the samples. Before attempting a systematic experimental program for minimizing the hydrogen content of open-hearth steels, it was necessary to determine the variation in hydrogen in the open-hearth bath during the normal processing of a heat. The logs of ten heats are illustrated in Figs. 1 through 7. The number at each point on the graphs refers to the number of pintube samples analyzed for hydrogen. An increase in the hydrogen content of the liquid metal during the early period of refining is apparent in all heats. There is evidence of a decrease in hydrogen later in the refining period, see Fig. 2, for example. Therefore, the existence of a maximum hydrogen content at some intermediate time during the refining period is anticipated. To test for the existence of a maximum, the data illustrated in Figs. 1 to 5 were grouped together so that they could be examined on a collectively comparable basis. The grouping was effected, as illustrated in Fig. 8, by superimposing Figs. 1 to 5 so that the apparent hydrogen maxima coincided at a fixed position, indicated by the zero on the time axis. Also, by vertical adjustment until the graphs appeared to the eye to be superimposed, the hydrogen scale was made to show the change in hydrogen content. That is, a standard amount was subtracted from the hydrogen values for each heat. Therefore, Fig. 8 represents a generalized presentation of the data and, for this reason, cannot be used to predict the actual hydrogen content of the bath during refining, but rather to show the probable rate of change in hydrogen during refining. Statistical examination of the grouped data in Fi.g. 8 shows that the apparent. increase and suhsequent decrease in hydrogen are real and confirms the existence of a maximum in general. The average increase In hydrogen during the early period of refining is between 0.8 and 1.3 ppm per hr, and the

Jan 1, 1958

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: 1-maintenance of metallurgical balance, 2-conservation of oxide ore, and 3-mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak-Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pet, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pet of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1952

By L. M. Yokum

Tremendous strides have been made in front end loaders in the last 15 years. Fifteen years ago a 21/4-cu-yd front end loader was the largest standard machine you could purchase. Today there are two commercially available 15-yd front end loaders and many 10-yd loaders. The estimated total worldwide market for all front end loaders over 1 cu yd has increased to over 30,000 units yearly. The trend for front end loaders, in general, and for the mining market, in particular, is to larger and larger loaders. Today there are 18 and 24-cu-yd front end loaders in prototype form. It is not inconceivable that within another 15 years there will be front end loaders in the 50-cu-yd class with design features quite different from today's. These larger loaders are being designed at the demand of the mining industry, and the design criteria are being directed at this market area. The costs of these larger loaders may be greater per unit of capacity, but it is the intent of the various manufacturers that they will be more reliable, be easier to service, give greater availability, live longer, and cost less to maintain, than previous loaders. It is also the intent of the manufacturers to compete directly with the large mining shovel as the front end loaders become appropriately larger and more reliable. There has been some trend in this direction in the past with smaller shovels and loaders. In Table 1 are data for U.S. production of front end loaders (5 to 6-cu-yd class and 10-cu-yd class) and U.S. production of cable-operated, crawler-mounted power shovels, draglines, and cranes (11/z to 5-cu-yd class) for the years 1965 and 1970 (Source: Bureau of Census, U.S. Department of Commerce). Many believe that the decline in shovel production, in particular, is due at least in part to the acceptance of the front end loader as a primary production tool. However, the foregoing comparison is for a smaller shovel class than for front end loaders. It has generally been a rule of thumb that a front end loader is competitive on a cost and production basis with a shovel when it is approximately twice as large as the shovel. This same relationship may not hold true with the larger 10 to 15-cu-yd electric shovels, but at some point, direct competition will probably occur. Current Trends There are many trends today that are significant. One major usage trend in large front end loaders involves load and carry operations. This application is defined as "performing the functions of loading, hauling and dumping with one machine over distances within its economical limits." Many people are using front end loaders to transport material as well as load it. It is estimated that nearly half of all 10-cu-yd front end loaders sold in 1970 went into primary load and carry work. Many times, significant savings result from using the load and carry system, instead of small truck and shovel operations, for distances in excess of 1000 ft. Strip mining in coal utilizes great numbers of front end loaders in this manner. In-pit crushing of limestone is developing on an ever-increasing level, especially in Europe, and this stimulates interest in load and carry operations. In Table 2 is a cost example for 10-cu-yd front end loader in a limestone quarry. More than likely, load and carry use will continue to increase over the next few years. While some types of mining do not lend themselves to load and carry in the pit, such areas as surge piles are applicable and should be studied. In the area of machine operating cost reduction, there

Jan 1, 1974

By D. J. DeLazaro, D. W. Levinson

Observations were made of the isothermal transformation and quench and reheat transformation characteristics of binary titanium alloys containing nominally 6 wt pct and 10 wt pct Mn at temperatures of 700°, 600°, 500°, and 400°C. A large change in electrical resistivity attends the beta to alpha transformation in these alloys. The correlation between the resistivity and microstructure provides a very sensitive means of following the transformation. IN a recent paper, Wyatt' determined the electrical resistivity of several grades of unalloyed titanium metal as a function of temperature. In addition to the expected discontinuity accompanying the a-B transformation, it was shown that s-Ti has a much larger temperature coefficient of resistance than 8. The result would be a very large change in resistivity attending ß?a: transformation if it were possible to retain the ß structure at room temperature in pure titanium. From these data, it was reasoned that this very large resistance change should accompany the decomposition of the ß phase in alloys in which the ß phase can be retained by quenching to room temperature from the ß field. This effect is illustrated schematically in Fig. 1. Binary Ti-Mn alloys containing more than 5 wt pct Mn can be quenched to retain ß from the ß field. While the ß phase would be expected to transform in these alloys to a + TiMn via a eutectoid reaction,' it has been observed that the eutectoid decomposition does not occur readily in hypoeutectoid alloys, but rather the decomposition of the ß occurs by rejection of proeutectoid a only." Thus, the resistivity should show a marked decrease, as the a: phase is rejected from ß and a study of the resistivity of solution treated and isothermally transformed or quenched and reheated alloys should provide an excellent means of establishing the transformation kinetics of these alloys at various temperatures. The usual metallographic method of determining the transformation temperature-time characteristics of alloys of this type is very satisfactory for deter- mining the time required to initiate the rejection of the a phase at temperatures relatively close to the ß/a: + ß boundary when the rejected a is coarse, and a typical Widmanstaetten structure is produced. The metallographic method is less satisfactory for this purpose at lower temperatures where a fine nodular decomposition product is observed and is very unsatisfactory for the determination of the time required to complete the transformation at any temperature. Accordingly, specimens were prepared for resistivity measurements in order to determine the transformation characteristics of the alloys.

Jan 1, 1955

By S. R. B. Cooke, H. S. Choi, I. Iwasaki

Oleic acid is the chief ingredient of fatty acids used as collectors in nonsulfide flotation. With a few notable exceptions, the various quantities of saturated and other unsaturated acids comprising the remainder are usually unspecified. There have been occasional investigations to determine the collecting ability of these other acids and their effect on collection with oleic acid, but much remains to be done in this field. Hukki and Vartiainen, on the basis of tests using palmitic, oleic, and linoleic acids and an impure linolenic acid, concluded: "The collecting power of fatty acids used in flotation increases with increasing unsaturation of the hydrocarbon chain." ' Their work was concerned mainly with the response of ilmenite to the foregoing collectors, although tests were also made with magnetite, pyrite, hematite, and rutile. Cooke and Nummela, employing an almost identical technical procedure to evaluate flotation of hematite, magnetite, and goethite with various fatty acids, including oleic, linoleic, and linolenic acids of the highest purity obtainable, found: "The effectiveness of 18-carbon unsaturated fatty acids, as collectors for hematite, decreases with increase in the number of non-conjugated double bonds." 2 Vn the foregoing investigations, collector efficacy was measured on the basis of reagent addition in moles per liter. Sources of fatty acids which may be used as collectors are legion, but provided the oleic acid content is high enough, cost is normally the determining factor. Usually this restricts severely the purity of the collector in terms of its oleic acid content. Within the past year or so users of crude oleic acid have become somewhat more discriminating, apparently realizing that some of the organic acid impurities may exert deleterious effects upon certain separations. Marine fish oils have been used as collecting agents, particularly in the earlier days of froth flotation. Currently their use is restricted in this field to relatively minor applications, although they are cheap sources of fatty acids. In an effort to increase the range of uses for fish oils, the U. S. Fish and Wildlife Service has awarded contracts to study these materials as sources of new products suitable for commercial development. Funds for these investigations were provided by the Saltonstall-Ken-nedy Act of 1954. One such contract was awarded to the University of Minnesota School of Mines and Metallurgy for the purpose of investigating the use of fish oil fatty acids and their derivatives in ore flotation. Besides the esters of various saturated aliphatic acids and of palmitoleic and oleic acids, fish oil also contains esters of monoethenoid and polyethenoid acids with 20, 22, and 2 carbon atoms. The basic purpose of the investigation was to determine whether the longer-chained unsaturated acids possess collective properties superior to those of the C,, monoethenoid oleic acid, the C16 monoethenoid palmitoleic acid, or the saturated C,, stearic acid and its shorter chain homologues; whether highly unsaturated acids have merit as collectors; whether rela-

Jan 1, 1960

By J. W. Whalen

The design, operation and evaluation of an instrumental method for the determination of fluid saturation in porous materials during multiphase flow studies is described. The presence of a magnetic tracer comprising approximately 20 per cent by weight of the aqueous phase is necessary in the utilization of this method. The results of two-phase relative permeability studies on two samples of porous media are presented. The realizable accuracy of the magnetic susceptibility apparatus in fluid saturation determination is estimated to be 2 saturation per cent. The response characteristics of the instrument, together with its simplicity of operation and maintenance are sufficient to recommend its use in routine laboratory studies. INTRODUCTION The problem of fluid saturation determination in laboratory multiphase flow studies characterizing oil recovery processes is one which has been given exhaustive study. Briefly the requirements which are considered essential for a method sewing this purpose are: (1) that the measurement be made external to the porous rock sample under study with no interruption of the fluid flow pattern, (2) the saturation indication must be independent of fluid distribution in the volume scanned by the measuring device, (3) the sensing element should have a small field of definition enabling several independent measurements to be made along the length of a core sample, and (4) the properties of the fluid phases should not be too greatly influenced by required tracer substances. Several methods of saturation determination have found acceptance as meeting most of the above requirements. These include resistivity,1 X-Ray absorption,2,3,4 radiotracers,5,6,7 and neutron8 diffraction The merits of these techniques have been discussed elsewhere in detail and will not be considered here. It may, however. be pointed out that the elaborate instrumentation required by the methods mentioned above, with the exception of the resistivity method (which is undesirable from other standpoints) encourages the development of methods of saturation determination which, while meeting the requirements listed, are simple to operate and maintain. It is the purpose of this paper to describe the instrumentation and evaluation of a technique for the determination of fluid saturation which combines relative simplicity with an accuracy and applicability comparable with methods now in practice. DESCRIPTION OF THE METHOD If the magnetic susceptibility of one of the fluid phases contained within a porous material is sufficiently greater than that of the remaining phases, the fluid content of the material with respect to that phase may be determined by an induction measurement. This may be accomplished by measuring the induced voltage in the secondary of a transformer which contains the porous sample and its interstitial fluids as a portion of the magnetic flux path. Since the fluids used in multiphase flow studies involving petroleum reservoir samples possess uniformly low magnetic susceptibilities, the use of a magnetic tracer in one of the fluid phases is required. In view of the relatively high paramagnetic susceptibility of salts of the transition elements, together with their high water solubility. these compounds become logical choices as tracer materials. Cobaltous chloride, having a specific magnetic susceptibility of 17.5 x 106 cgs electromagnetic units, was used as the tracer material throughout this work, in a concentration of approximately 20 per cent by weight in the aqueous phase. MAGNETIC SUSCEPTIBILITY APPARATUS Basically, the magnetic susceptibility apparatus consists of a transformer serving as the detecting head, a

Jan 1, 1955

By R. C. Buehl, M. B. Royer

A three ton per day blast furnace using blast temperatures up to 2200°F was operated to recover manganese from open-hearth slag and manganiferous iron ore. The spiegel product containing 12 to 24 pct Mn was blown in a basic converter to high manganese oxide slag as described in the paper which follows this one. HE steel industry, which uses over 900,000 short tons of contained manganese yearly, depends on the maintenance of supplies of manganese from foreign sources such as India, Gold Coast, and Union of South Africa. Domestic ores furnish less than 10 pct of the yearly requirements of metallurgical ores and, unfortunately, the United States has no large, or even moderate, grade deposits that could supply a major portion of the requirements for many years. It may be necessary to obtain manganese from low grade materials, even though the cost is appreciably higher than when imported ores are used. An American Iron and Steel Institute questionnaire prepared in 1949 revealed the startling fact that open-hearth slags contained about as much manganese as is imported annually. The quantity of slag produced each year by integrated steel plants (not including cold-melt shops) and the manganese contained in the slag follows: Flush slag, 4,500,000 net tons a year containing 408,000 net tons Mn; tap slag, 7,230,000 net tons a year containing 364,000 net tons Mn; for a total of 772,000 net tons Mn. Some of this slag is returned to blast furnaces but about 75 pct would be available for the recovery of manganese. In several decades, when many of the present sources of high manganese ores have been depleted, open-hearth slag and manganiferous iron ore may become the major source of the manganese required for steel production. In 1948, manganese-ore imports from the U.S.S.R. were 427,000 net tons or 34 pct of the foreign ore. Imports from this source were only 81,000 tons in 1949. It became evident that adequate stockpiles of manganese could not be accumulated for many years. The Bureau of Mines initiated an extensive research program on methods for increasing domestic production of manganese and entered into a cooperative agreement with the American Iron and Steel Institute for recovering manganese from open-hearth slags whereby the Institute provided additional research funds and trained operating personnel. Description of Process and Equipment Numerous pyrometallurgical and hydrometallur-gical processes and combinations of the two can be applied to the recovery of manganese from open-hearth slag and manganiferous iron ore. The process being developed at the Pittsburgh Station of the Bureau of Mines consists of: 1—Reduction of open-hearth slag in a blast furnace to recover the iron and manganese as spiegel-eisen with 12 to 24 pct Mn. 2—Selective oxidation of the manganese from the blast-furnace metal in a basic-lined converter to produce a high manganese slag with up to 60 pct Mn content, which is a synthetic ore for the production of ferromanganese. This process has the advantage over other processes in that the equipment which would be used, such as blast furnaces and converters, has been highly

Jan 1, 1953

By T. Leffers, A. Grum-Jensen

The development of texture in copper and brass (15 pct Zn by weight) rolled at room temperature and in copper rolled at -196°C has been followed by determination of pole figures for various degrees of reduction. The concurrent development in microstructure has been investigated by transmission electron microscopy. For low deformations, the texture in brass, and in copper rolled at -196°C, is the same as the texture in copper rolled at room temperature. The difference in texture does not show until 50 pct reduction. At the same reduction at which the difference in texture appears, mechanical twinning starts in brass and low-temperature rolled copper. Nevertheless, Wassermann's twinning theory is found to be incompatible with the present resulls. All fcc metals and alloys develop, by heavy rolling, one or the other of two types of texture, classified for convenience as copper type and brass type, or possibly a '(mixed" texture representing an intermediate stage between the two types. The pure fcc metals, except silver, develop copper-type texture by rolling at room temperature. Silver and many alloys, including brass with more than 10 pct Zn, develop brass-type texture.' By rolling at low temperatures a transition in texture from copper type to brass type has been observed for copper of commercial purity.2'3 It is generally accepted that there is a correlation between the type of texture developed and the stacking fault energy. Metals and alloys having fcc structure develop brass-type texture when rolled at room temperature if the stacking fault energy is low, whereas those with higher stacking fault energies have copper-type texture.2 Thus, the difference in stacking fault energy must cause a difference in deformation mechanism responsible for the difference in texture. Of the numerous theories advanced concerning the nature of this difference in deformation mechanism, four seem to be of current interest. Smallman and Greeen,2 Dillamore and Roberts,4 and Dillamore5 refer to the fact that cross slip more easily takes place in a material with a higher stacking fault energy and suggest that the copper-type texture is developed by deformation with extensive cross slip whereas the brass-type texture is developed by deformation without cross slip. Haessner6 has advanced the theory that the copper-tye texture is developed by a combination of normal {111} slip and slip on other planes, mainly 1100). Thus his theory, like the theory above, proposes an additional deformation mode for the formation of the copper-type texture. Contrary to this, wassermann7 and Ahlborn et al.8 believe that the copper-type texture is obtained by normal slip. In their opinion mechanical twins will change a copper texture into a brass texture. For this reason it is suggested that the brass-type texture is developed when, in addition to normal {111}(110) slip, mechanical twinning is present. None of these theories are, however, supported by direct experimental evidence. In a recent experiment Hu et al.9 have followed the development of rolling texture in copper and brass. The texture of copper and brass was found to be es-

Jan 1, 1969

By W. L. Phillips

The plastic properties of lithium fluoride and magnesium oxide under compression were investigated in the temperature range 25° to 1000°C. At the higher test temperatures, the critical resolved shear stress and the rate of work-hardening of both materials decreased. At the same time, the number of stress lines decreased and the individual stress lines became wider. The density of dislocations decreased for a given strain the higher the test temperature for lithium fluoride. The decrease in the critical resolved shear stress as a function of temperature could be fitted by the equation 7 = Toe-KT. THE influence of temperature on the stress-strain curves of a variety of metal single crystals has been studied for a wide range of temperature.1"5 In general, it has been found that both the critical resolved shear stress and the rate of work-hardening decrease rapidly with increasing temperature and fall to a value of essentially zero well before the melting point is reached. AS the test temperature is increased, the density of slip lines has been found to decrease, while the width of the individual lines increases. Recently, Gallagher,' Graf et al.,7 and Pate1 and Alexander8 have found that single crystals of germanium and silicon, which have a covalent bond, can be plastically deformed at elevated temperatures. In compressing germanium it has been found that the strain-hardening coefficient decreased very rapidly as the temperature was increased. The coefficient did not tend to zero, however, as the melting point was approached. The plastic properties of ionic crystals at elevated temperatures have not been extensively investigated. In the only work previously reported Tamman and salge9 and Theile10 have shown that the yield point of sodium chloride single crystals tested in tension and compression decreased slowly with increasing temperature, falling to zero at the melting point. As the temperature was increased, they found a substantial increase in the amount of plastic deformation that could be sustained before fracture; and at temperatures above 400°C in tension tests there was an obvious necking which produced local extensions of up to 3000 pct. The purpose of the present work is to extend the knowledge of high-temperature deformation to the ionic crystals, lithium fluoride and magnesium oxide. The following were investigated: 1) The gross shape of the stress-strain curve at several temperatures. 2) The change in the critical resolved shear stress as a function of temperature. 3) The appearance of slip bands and the density of dislocations as a function of strain and temperature. EXPERIMENTAL TECHNIQUES The lithium fluoride and magnesium oxide used in this investigation were purchased from the Harshaw

Jan 1, 1961

By Elizabeth L. Robinson

INTRODUCTION The financial structure of the mining industry has changed strikingly over the last thirty years. The changes that have taken place in the last ten years include both the acceleration of adverse trends in familiar balance sheet numbers and ratios, and the subtle but extensive increase in the complexity and diversity of funding sources, both on and off the balance sheet. The objectives of this paper are: 1. to briefly trace the evolution of trends in the mining industry's financial structure; 2. to focus in more detail on changes in the last decade; and 3. to analyze those trends both within the framework of capital structure theory, and in the context of changes in business and economic conditions. This survey focuses primarily on the mining companies listed on the major U.S. stock exchanges, since financial data for these firms are readily available. The producers of minerals are a diverse lot, however, and the quantitative data on "the mineral industry" as characterized in this survey fails to encompass the financial arrangements for the bulk of production worldwide, whether owned by governments or private-sector enterprises. Even large portions of the U.S. minerals industry - subsidiaries of still larger firms, medium-sized privately-held companies, and the small miners - were not included in the quantitative part of the analysis. THE EVOLUTION OF THE INDUSTRY'S FINANCIAL STRUCTURE In the aftermath of the business failures of the Great Depression, most borrowers and lenders considered minimal debt on the balance sheet the soundest policy. By the 1950's, however, a body of literature was developing which demonstrated that the use of leverage could improve the value of the firm, given a corporate tax structure which allows interest, but not dividends, to be deducted from income. Optimal Capital Structure The objective of undertaking a new project is to improve the wealth of the current owners of the firm. If internally-generated funds are insufficient to allow growth at the desired rate, the company can raise funds either by issuing new equity, or by incurring debt. The most attractive source is the one which will provide the funds for the lowest cost. Both stockholders and lenders expect to receive compensation for use of their funds at a rate which reflects two things: (1) the return they could earn in alternative investments, and (2) the riskiness of the projected return. Factors which increase business risk include the variability of demand, sales prices and input prices, and "operating leverage", or the proportion of fixed to variable costs. In addition, from the investors' point of view, the use of financial leverage concentrates risk onto the shareholders, because lenders enjoy a prior claim in the event of bankruptcy. So although financial leverage can increase a project's expected rate of return, and therefore, presumably, earnings per share, too much debt will be perceived as uncomfortably risky by equity investors, and have a negative effect on stock prices. In theory, then, there should be an optimal level of debt which the firm's managers should seek to obtain in its financial structure. The desirable proportion of debt should in general be lower for a firm with a higher inherent business risk. In addition to their own judgement, financial managers are sensitive to the attitudes of lenders, rating agencies, and investment bankers in setting financial targets (Scott and Johnson). In practice, financial structure will also depend upon the company's growth rate and level of profitability, and conditions in the capital markets at a given time.

Jan 1, 1985

By C. B. Scotty, E. F. Egan

INTRODUCTION The neutron-garnma log has been used for stratigraphic correlation by the oil industry for a number of years. In the past few years, the quantitative application of the log to provide information with respect to porosity has received considerable attention. The basic concept of quantitative interpretation has been pointed out in the literature.1 It shall be the purpose of this paper to develop by empirical means a method whereby the relationship of porosity to neu. tron-gamma deflection may be predicted for various bore hole diameters. This relationship is applicable to the present commercially available neutron-gamma log run in open hole. Data available at present are insufficient to establish similar correlations for cased holes. Through the development of such a correlation, the importance of the various factors which influence the log become evident. NEUTRON-GAMMA CURVE The logging tool conisists of a source of neutrons and a means of detecting and measuring gamma rays, the detector being shielded from the direct emission from the source. The gamma rays reaching the detector have three sources of origin: (1) the natural gamma radiation emitted by the formations, (211 induced gamma radiation resulting from neutron bombardment. (3) gamma rays emitted by the source and scattered by the formation. It has been found that a neutron derived porosity correlation is best for shale-free formations. Under this condition the natural gamma ray radiation is quite low, and therefore, will be neglected in this work. The induced gamma ray radiation resulting from neutron bombardment is predominantly a function of the hydrogen atom concentration of the substance penetrated. Due to the geometry of the presently used system, the neutron-gamma curve indicates low radioactivity opposite high hydrogen concentration and high radioactivity opposite low hydrogen concentration (Figs. 1 and 2). Having a curve which measures hydrogen atom concentration, it follows that if the matrix of the formation penetrated consists only of atoms other than hydrogen, and if the fluid filling the pores is either oil or water, the curve should be an indication of porosity (both oil and water have practically the same number of hydrogen atoms per unit volume'). The above criteria for a porosity (Correlation are fulfilled in clean sandztones and limestones in which the pore volume is completely filled with liquids (oil or water), but not in formations containing either free gas or shale or both. Free gas. even under pressure, contains considerably less hydrogen per unit volume than oil and water; therefore, a variation in response on the curve would be expected. Camp- bell and Winter,' indicate that this difference in hydrogen concentration is sufficient to distinguish between sands containing water or oil and those containing dry gas. Shale is a mineral that contains hydrogen in its molecular structure in addition to the hydrogen which is present in the water of hydration. Thus, in formations containing variable amounts of shale, the hydrogen concentration indicated by the curve is not a direct indication of porosity. Bush and Mardock' have indicated that a correction may be made by use of the gamma ray curve to provide a porosity correlation. Since the induced gamma rays are predominaiitly a function of hydrogen concentration, it would be anticipated that the fluid filled bore hole would materially influence the logs. It has been pointed out that the neutron-gamma curve response decreases with increase in hydrogen concentration, therefore, the response due to the induced gamma ray component decreases with increase in fluid filled hore hole diameter. The scattered gamma ray component of the log increases with increase in bore hole diameter. Although there is an

Jan 1, 1952

By B. N. Webber

During World War II the Pocos de Caldas bauxite deposits of Minas Gerais in Brazil yielded some 60,000 tons. Since then they have maintained a small but almost continuous production. Known for many years, the deposits have been reconnoitered by a number of geologists,1-4 but no dimensional study has been possible because there have not been enough exposures to show clearly the depth, bedrock, and habit of occurrence. The deposits are somewhat unorthodox in their occurrence mainly on steep slopes of mountainous terrain. Although the parent rocks are a sequence of alkalic intrusions that bauxitize readily under the proper conditions, bauxitization has not been the dominant weathering process in the area. It is limited to certain localities where important volumes of bauxite have formed by laterization of several rock types, each yielding a distinctive type of bauxite. Bauxitization may be observed in several stages and in both direct and indirect cycles of formation. The deposits are unique in being of demonstrably recent age, so that the conditions under which the bauxite formed are essentially those of the present time. The bauxitization sequence is postulated in Fig. 1. Pocos de Caldas district is situated at approximately 21" 50' south latitude and 46" 30' west longitude in southwestern Minas Gerais, near its boundary with the state of Sao Paulo. It lies within the Eastern Plateau physiographic province of Brazil, occupying a prominent highland area that rises some 2000 ft above the surrounding country. The central portion of the Pocos de Caldas highland is characterized by gently undulating topography, with generally sluggish drainage, and is surrounded by a rim of hills rising above 1000 ft above it. In cross section the highland resembles a truncated cone with a depressed central area. The rim has been sculptured into a more prominent and abrupt topographic pattern than the central area, and streams that are of low to moderate gradient in the central plateau plunge over the rim in a spectacular series of cascades and falls. GEOLOGIC SETTING The Pocos de Caldas highland owes its physiographic expression mainly to the differential erosion of a complex intrusion of alkalic rocks into gneiss and sandstone. Erosion was guided somewhat by faulting. The highland is almost circular, and its limits, except in one locality, coincide with the periphery of the intrusion. The alkalic intrusions largely control its shape and extent, aided by some subsidence through faulting of the central portion. The mountainous rim of chilled phase, the later intrusions, and the indurated intruded rocks have resisted erosion more effectively than the central portion of the intrusion, which is composed of more coarsely crystalline alkalic rocks that form a central basin in the highland. Gneiss and sandstone are found coincident in elevation with the topographically highest intrusions. BAUXITE TYPES Most of the bauxite, which varies in physical type, is on the top and steep slopes of the northern rim. The only alumina mineral is gibbsite, definitely crystalline but iron-stained unless of secondary origin. Clay particles occurring with the bauxite are kaolinite. Some types of bauxite are an index to their origin. There are four general physical types of bauxite or bauxite laterite: 1) Surface Rubble: This partly nodular and partly fragmented type, usually less than 3 ft thick, results from surface weathering of massive bauxite or may form by superficial enrichment of nodular bauxite with a clay matrix when the interstitial clay is removed by surface runoff.

Jan 1, 1960

By Peter A. Allen

INTRODUCTION Lac Minerals Ltd., a significant North American producer of gold and other metals for over three decades, recently completed the financing for its third discovery in an eight year period. From 1975 to 1982, three important gold deposits in Quebec and Ontario: namely Bousquet, Doyon and Hemlo, were located and tested by company geologists. When current development plans are complete, Lac will have injected $400 million into these deposits. A discussion of these projects and their financing reveals some extraordinary challenges faced by top mining management. Responses to these challenges developed by Lac management may interest executives planning to proceed with new North American gold or base metal mines. MINE DE BOUSQUET In 1976, we explored a property containing a large low grade surface gold deposit. The geologial setting was exciting and we tested recent rethinking about the volcanigenic genesis of gold and base metal occurrences. These ideas helped pus to persist in drilling a nearby conductor and 600,000 ton, 0.22 ounce/ton orebody was outlined. With gold at $Cdn.150 per ounce the feasibility was approved and work started February 1978. Availability of a company custom mill 35 miles away reduced our immediate capital needs to $8.5 million. First of all, understanding the geological significance of results at Bousquet was critical. Mine and surface exploration geologists at Lac are continually and closely working together to better understand the nature of gold deposits in the Shield. This management emphasis on both pure and mine geology meant that a much larger ore reserve could confidently be predicted. (A proposal to quickly develop and ship the ore to a nearby company mine was discarded.) Instead of designing just for the initial plan of 500 tons per day we therefore included headframe, shaft and hoisting capacity for 1,700 tons and provided a compartment for future shaft deepening. These prebuilt additions were welcomed within two years. The type of collaboration whereby surface and mine geologists make study and field projects of each other's problems sets the foundation for this predictive ability. At Lac, the traditional responsibility for and supervision of the mine geologist by the mine manager is modified by this collaboration. This adds interest and scope to the geologist's work and lets him think beyond production requirements. Besides such management considerations, dedicated, professional geologists with integrity and lots of practical experience are needed to plan successful developments like Bousquet. Second, projecting future bullion revenue was critical and our experience selling production for future delivery gave us a way to ensure that loan repayment at term would not be jeopardized by falling metal prices. Output from a single mine is often reduced or curtailed for numerous reasons, so "selling forward" is done best and with least risk by companies like Lac with more than one or two producing locations for that metal. At first, a forward marketing program was considered unwise, speculative and unorthodox by analysts and other producers. Declining markets in the early 1980s prolonged gold project capital recoveries however, and critics were better able to appreciate the usefulness of such a tool. The notion that gold miners shouldn't sell forward is rooted in the belief that long-run gold prices will rise as modern social theory implementation in democracies inevitably erodes currency values. Selling against this trend goes against this industry belief or "ethic" and is at odds with the company shareholder who expects the gold price to rise not fall. Nonetheless, an efficient tool is available to the gold producer (not to most other metal producers) to quickly disengage from adverse price trends or to lock in spot prices instantaneously for two years or more of profits. Gold had recently retreated from U.S.$200 to $104 and now had reversed to $150. Lac had been forced to close several of its producers during the late 1960s because central bank-controlled

Jan 1, 1985

By James M. Schopf

Gilbert H. Cady (Senior Geologist, Head of Coal Division Emeritus, Illinois State Geological Survey, Ur-bans)— Those coal mining and preparation engineers and operators who read this article will probably be most concerned with those parts which deal with the applied aspects of coal petrology in the fields of mining and technology. Actually the two categories overlap to such a degree that they are essentially one. The article appears to suggest that there are two points of veiw in the field of applied coal petrology. From one viewpoint the interest is in the attributes of the various macerals (Fig. 4)—vitrinite, exinite, and inertinite—-in terms of utilization (when the coal is burned, coked, gasified, hydrogenated, etc.). From the other point of view, assuming that the macerals do possess individuality in behavior, interest attaches to the possibility that coal of a particular type (maceral composition) can be prepared by discriminating methods of preparation and blending. Most of the 29 items listed for mining and technology can be assigned to one or the other of these two categories. Somewhat the same simplicity that resides in the items of the proximate chemical analysis is desirable for the petrologic analysis. This simplicity appears to be fairly well achieved by the group maceral device provided mineral matter is also included. At present, without fully adequate experimental substantiation, the three types of material represented by the group macerals vitrinite, exinite, and inertinite are regarded as sufficiently different in composition and properties so that the preponderance of any one type gives the coal distinctive properties provided it is not too high in rank. This appears to be increasingly true with coals of decreasing rank from medium volatile bituminous coals downward. The article reflects the great emphasis placed on the technique of descriptive coal petrology in America. Such description has resulted in a good understanding of the nature of the fundamental variability in the physical composition of bituminous coals. Such understanding has been attained very largely by the thin section technique used in coal bed profile analysis introduced by R. Thiessen after the manner suggested by the data assembled in Table 11. In contrast, the group maceral concept of coal analysis in terms of the categories indicated on the left hand side of Fig. 4 is particularly useful for broken coal analysis and applied coal petrology, in which evaluation is largely based on quantitative concepts. It is inadequately emphasized in this article that although the Thiessen procedure is suitable for the purpose of petrologic description in which texture and orientation are of critical importance, it is less well adapted to the technological uses of applied coal petrology. No doubt those persons whose interest may have been aroused in the possibilities of more refined coal preparation by this and other recent contributions to the literature of coal petrology will understand that certain technical aspects of coal petrology, particularly in the field of descriptive petrology, are still a matter of some controversy among the coal petrologists, but they will also appreciate that Schopf has made an important contribution to the field of coal petrology in America by pointing out that such petrology requires consideration of disoriented or broken coal samples and that results are expressed less in descriptive terms than in quantitative values with respect to the four types of material composing coal—-vitrinite, exinite, inertinite, and mineral matter. Unfortunately because so much American coal petrology has been descriptive of the coal in the bed there is a dearth of literature on applied coal petrology. Research and experimental data in this field are greatly needed.

Jan 1, 1959