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By Albert Sauveur

Some writers believe that when austenite transforms completely into pearlite on slow cooling through the thermal critical range, such transformation does not imply the formation of any of the so-called transition constituents: martensite, troostite and sorbite. Might it not be claimed with as much justification that when water freezes it does so suddenly, solid ice being obtained without passing through the intermediate stages of ice and water? Obviously, regardless of the speed of solidification, this would be erroneous. Before reaching the completed solid condition, this must of necessity pass through an unbroken series of stages when the system consists of mixtures of water and of ice, the former decreasing and the latter increasing in amount. Are not these "transitionV stages in the freezing of water similar to the transition stages occurring in the transformation of austenite into pearlite? To clarify, let us consider in Fig. 1 the mechanism of the transformation of eutectoid steel from its austenitic to its pearlitic condition as it cools slowly through its thermal critical range. It is generally admitted that this transformation is the result of two distinct phenomena: (1) the allotropic change of gamma iron into alpha iron and (2) the crystallizing out of solution of the carbide Fe3C. Whether this crystallizing of Fe3C implies also its formation is immaterial for our purpose. In other words, whether it is the carbon in atomic dispersion which is present in gamma iron or molecules of Fe3C has no bearing upon the matter to be discussed. The transformation of gamma iron into alpha iron cannot take place suddenly. It must of necessity be progressive, regardless of the speed with which it may take place. At A, for instance, there would be present AB gamma iron and BC alpha iron. The distance MN does not refer to falling temperature, but to the passage of time. It may be assumed that the transformation takes place at a constant temperature (the Ar321 point). In a similar way, the crystallization of Fe3C out of the austenite solid solution must be progressive as depicted between M and N. At A there would be present AB gamma iron, BC alpha iron, CD carbon dissolved in gamma and in alpha iron and DE Fe3C. This condition may be described according to individual preference as martensite or troostite. To sum up, can the metal pass from condition

Jan 1, 1931

By A. D. Brokaw

The so-called "paraffin dirt" of the Gulf Coast oil fields has been considered an indication of the possible presence of oil and gas, and not a few wells have been brought in solely on the basis of such evidence. In a sense, it may be empirically justified as evidence of the presence of oil or gas, though it cannot be said to be infallible. If the writer is correctly informed, its association with oil and gas was pointed out by Lee Hager some years ago, and while it is commonly mentioned by geologists and operators in conversation, no discussion seems to have found its way into print. The present interpretation is based on laboratory study rather than on field investigation, and is put forth with the hope of stimulating discussion which shall lead to more complete understanding. The term "paraffin dirt" has been applied to soils with a peculiar texture, which has been described as "curdy" or "rubbery." When moist, the material breaks much after the fashion of "green" cheese. It is rubbery under compression, but does not resemble rubber in tenacity or cohesion. When dry, the material ranges from hard clods to a horny mass of high tensile strength. The shrinkage in drying is very great, especially in the hornlike matter. The moist material ranges in color from dark brown in the specimens rich in organic matter to grayish in specimens containing more inorganic matter. All of the specimens studied were readily attacked by molds, which in some cases grew in great profusion on the samples. Samples exhibited a rather characteristic "swampy" or "mucky" odor when wet, and a characteristic odor of humic soils when moist. If the material is brought into suspension in water, very fine sand settles out and may be recovered by successive washing and decantation. The sand is almost entirely quartz, but contains occasional fragments of feldspar—more or less kaolinized and iron-stained. Practically all of the sand passes through a 100-mesh screen and the grains are extremely sharp —possibly below the limit of rounding. A flocculent mud, of clay and organic matter, remains in suspension for days. Some material is taken into solution in the water, imparting a slight tendency to froth when vigorously shaken. The results of partial analyses of four samples of "paraffin dirt," and, for comparison, one of a coastal plain soil are given in Table I.

Jan 1, 1920

By -1ng. B. Stampa

Introduction The author investigates the technology available for mine flow reversibility by reversing the operation of the main fans. Air bypassing, reversed sense of impeller rotation, and blade angle overturning are indentified as the most important means of achieving the flow reversibility. The axial flow fan is studied in some detail and the important variables of fan capacity, pressure, and efficiency are compared. Although some references to rotor velocity vectors are made (Fig. 1) the subject is not presented in detail in order to stay within the paper's intent. Background Where applicable, the requirements of mine air reversibility are an important part of the overall mining operation and should be given proper attention during mine ventilation design. The need for preventing mine fires in intake shafts is the reason that, for a long time, it has been a requirement in the Eastern European countries to have the feature of flow reversibility built into all main mine ventilators. In West German coal mines, this requirement does not apply since higher priority is given to safety precautions. Some metal and mineral mines in the United States are required to have ventilation currents readily reversible in order to fight underground fires as stipulated by the Title 30, Mineral Resources, and by the mining laws of some states. From an equipment performance standpoint, assuming the validity of other flow technical data, the question of the ventilator flow reversibility is of ten related to the actual scenario of the begining of an emergency. Most of the time, the system operates within half of the design capacity only and assuming that the equivalent mine orifice is stable only a quarter of the design pressure must be produced by the fan in any event. For such an operating point, the effective gradient and energy losses can be disregarded since fire emergencies usually last for only a short period of time. Equipment Comparison Fig. 2 depicts a ventilation arrangement wherein the flow can be reversed through changing the functional position of the dampers (slideplate or blade type). Both axial or radial fans can be applied in this case. For the mine exhaust operation, slide plates (guillotines) I and I11 are opened while plates I1 and IV are closed. The flow thus encounters two reroutings, which due to the reverse ventilation flow cannot be equipped with proper sheet metal deflectors, and therefore large pressure losses result during the air reroute operation. When the intake operation is considered, the dampers I and I11 are closed and the pair I1 and IV is opened. For this arrangement, almost the same flow conditions exist as for the foregoing exhaust mode; however, one additional corner must be applied. Even if a highly efficient ventilator is used in one of the previously described systems, the additional pressure losses resulting from air rerouting plus those encountered between the planes "0" and "00" significantly affect the overall energy requirements. Illustrated in Table 1 are the relative component losses of even flows for evaluation of various detour arrangements. From the aerodynamics standpoint, when these components are being designed, serious consideration must be given to the individual levels of the pressure loss coefficient, 5. Should there be an increase in the sum of the loss coefficients due to the particular component selection such that, for example, the net change is 1.2; the resulting pressure to be produced by the fan can be calculated from:

Jan 1, 1981

By J. S. Kasper, R. M. Waterstrat, B. F. Decker

THE existence of v phase in the Mn-Mo system was discovered in this laboratory several years ago,' but neither the conditions of its formation nor its composition were known, and preliminary attempts to reproduce the phase were unsuccessful. We wish now to specify the information appropriate to phase formation on the basis of experiments performed since that time. The starting materials were in every case electrolytic manganese and high purity molybdenum. Two methods of preparation were employed: l—arc-melting varying combinations of the elements in an inert atmosphere and subsequent heat treating, and 2—sintering compressed mixtures of finely divided powders of the elements. In the latter method, the powders were first passed through 200 mesh screen and then compressed into cylinders of % in. diameter and % in. length, which were sintered at elevated temperatures in an atmosphere of dry hydrogen. For some samples, the sintering temperature was 1200°C (for 1 hr), and these were then heat- treated at lower temperatures for times up to 24 hr. Eventually it was found, for the proper composition, that essentially pure a phase could be produced by sintering at 1125°C for 3 hr. Quenching to room temperature was effected by means of a stream of cold, dry hydrogen in a water-cooled chamber. The arc-melting procedure was found to be unsatisfactory after an investigation of many samples at different compositions and temperatures. Large quantities of manganese were vaporized during melting and none of the products were homogeneous or contained extended regions of pure phase. Nonetheless, these experiments established that no formation occurs in the Mn-Mo system below 1115oC, and indicated, at least roughly, the composition range of interest for further exploration. The more definite results obtained from samples produced by the second method are given in Table I. A portion of sample 2 heated for 16 hr at 1000°C in a Vycor tube transformed to molybdenum plus unidentified phases of complex X-ray pattern. From these few results it appears that the region of phase is quite narrow in composition. It is of interest that the composition is such as to make for difficulties in correlating this structure on an electron atom basis with other a structures of iron, cobalt, and nickel. In that respect Mn-Mo shows the same anomaly as Mn-Cr and Mn-V systems.' A second feature is that in common with the two other known molybdenum (Fe-Mo and Co-Mo), Mn-Mo v exists only at elevated temperatures. The X-ray powder pattern of the pure a phase of Mn-Mo is given in Table 11. It agrees well with the patterns established for other structures, especially with that for Fe-Mo as given by Goldschmidt.An attempt is being made to establish whether an ordering of the respective atoms occurs and such information will be reported in a future publication. Acknowledgment The authors acknowledge gratefully the assistance of Mrs. J. R. Belanger in some portions of this work. References 'J. S. Kasper, B. F. Decker, and J. R. Belanger: Journal of Applied Physics (1951) 22, p. 361. 2 A. H. Sully: Journal Inst. Metals (1951) 80, p. 173. H. J. Goldschmidt: Research (1949) 2, p. 343.

Jan 1, 1954

By R. B. Full

DURING the summer of 1949, the United Nations Scientific Conference on the Conservation and Utilization of Resources met at Lake Success. As summed up by one writer, the purpose was: "That everyone will try to learn how to feed more people from an acre of land." That objective is a stimulus to the phosphate rock industry as well as other industries. Phosphate rock was mined first in the Canadian province of Quebec where it was found as an apatite in pockets. In 1863, mining began near the Rideau Canal, first in open trenches and quarries until shafts were sunk. These operations were difficult since hard rock had to be drilled and blasted. Steam was the source of power. The phosphate rock was hand-selected and recovery ranged from 6 to 10 pct, averaging 70 pct to 85 pct bpl (bone phosphate of lime). Costs, f.O.b. Montreal, were $14 a ton, and selling prices averaged $17 a ton. In 1885, about 29,-000 tons were sold on this basis, but by 1892 the output fell to 8000 tons at an average price of $15 a ton at Montreal. South Carolina Deposits Phosphate rock was first discovered in South Carolina in 1837, but a mining company was not formed until 1867. It was immediately successful, and the industry continued to develop and ship an increased tonnage of land and river rock, shipments increasing from approximately 20,000 tons in 1868 to 537,000 tons in 1890, when the shipments began to decrease. In 1892 only 350,000 tons were shipped. In this field there were two types of deposits: land rock and river rock. More land rock tonnage was mined until about 1887 when river rock production began to equal it. The land rock was mined adjacent to and in the river marshes by pick, shovel, and wheelbarrow. Each worker dug daily a space 6 ft wide, 15 ft long and 6 ft deep, which was overburden, and then loaded about 3 cu yd of phosphate rock into wheelbarrows. His helper would wheel the barrows to small steam railroads up to 250 ft distant. Water was kept from the pits by steam pumps connected to railway cars. The phosphate rock was loaded into railroad cars, hauled and dumped into a log washer to which water was added, and run over trommel screens to eliminate the clay and loose sand. Steam shovels for loading were introduced about 1891. River rock was mined by floating dipper dredges and at low tide was loaded into barges by laborers using oyster tongs, the large lumps being hand-loaded. Barges and bateaus were unloaded by wheelbarrows which were wheeled to the same type of land rock washers. The washer product was piled with stacked pine cordwood in alternate layers and fired for drying. Usually this was done under sheds along the river bank. The dried rock then was transferred to ships, usually three or four-masted schooners. Cost of production of both land and river rock, including $1 a ton royalty paid to the state for state-owned property, was approximately $4.25 a ton averaging 60 pct bpl. The selling price was approximately $7 a ton f.O.b. In 1886 the courts handed down a decision permitting anyone to mine, and many additional companies were organized immediately. As a result, the selling price was lowered to a profit margin of 50Ø or less per ton. The South Carolina mines were eventually shut down when a more profitable source of phosphate rock was discovered in Florida. Florida Deposits Phosphate rock was discovered in Florida in 1882, and in 1883 a small quarry was opened near Hawthorne. In 1886 phosphate rock of good grade was discovered on Peace River, and the Arcadia Phosphate Co. was formed. In 1888 the company began shipments amounting to 3000 tons that year. During the latter part of 1888, hard-rock phosphate was discovered near Dunnellon. Exploration extended rapidly, and deposits of hard-rock phosphate were mined over a wide area so that shortly after 1900

Jan 1, 1952

By N. A. Warner

Dense cylindrical specimens of artificial hematite were reduced in hydrogen over a range 0-f total pressures between 0.1 and 1.0 atm and temperatures between 650" and 950°C. Hydrogen reduction at a total pressure of 1.0 atm in the presence of a chemically inert diluent and reduction in hydrogen/water mixtures were also studied Znter.face movement during reduction was followed using metallographic techniques. Substantiu1 partial-pressure gradients are shown to exist in both the gas film and the reduced iron layer. The results are consistent with a mixed-control mechanism involving an interaction of gaseous-difusion effects with a first-order reversible chemical reaction at the iron/wüstite interface. Over the hydrogen-pressure range of 0.1 to 1.0 atm, the rate is not directly proportional to the hydrogen pressure in the bulk gas stream. Nitrogen dilution of the reducing gas masks the true effect of hydrogen partial pressure by influencing the effective molecular dif-fusivity and gives an apparent first-order relationship. The transport of hydrogen and water vapor across the iron layer is consistent with molecular gaseous diffusion rather than a Knudsen diffusion mechanism. ALTHOUGH the literature on the reduction of iron oxides is very extensive, it is only comparatively recently that a definite quantitative model for the reduction of hematite has been proposed. This development has been largely due to the work of McKewan. An extensive literature review of previous work has recently been compiled by Themelis and Gauvin. Exclusive control by gaseous diffusion in the reduction of hematite has been recently advocated by Kawasaki and coworkers.' Although the results of the latter workers are of considerable practical significance, the extreme porosity (approximately 30 pct) of their specimens precludes their application to a kinetic evaluation of the reduction of dense polycrystalline hematite. The salient features of the kinetic model used by McKewan for the gaseous reduction of dense hematite at temperatures in excess of 560°C (when FeO is stable) are: 1) The reduction is a topochemical reaction taking place through 'the series FeO/FesO4/FeO/ Fe; 2) The gas-solid type of reaction takes place only at the FeO/Fe interface and the internal reduction: FeO proceeds by solid state diffusion; 3) The iron layer formed is quite porous and offers negligible impedance to gaseous diffusion; 4) The rate-controlling step in the reduction is a chemical process occurring at the Fe/FeO interface. In view of the relatively fast reaction rates involved when hematite is reduced in a stream of hydrogen, it seemed possible that the simplifying assumptions of negligible diffusional resistance in both the bulk gas phase and the reduced iron layer could introduce serious errors in the interpretation of experimental reduction data. The present work was initiated to elucidate the effect of gaseous diffusion on the reduction rate and to determine its possible influence in the formulation of the reaction mechanism at the metal/oxide interface. EXPERIMENTAL Procedure and Apparatus. Dense cylindrical compacts of artificial hematite were prepared by sintering pressed compacts of reagent-grade ferric oxide, containing 99.6 pct FezO, in air at 1200°C for 80 hr. The resulting compacts had a final porosity of less than 2 pct and a height and diameter of approximately 1 cm. Variations in sintering atmosphere from air to pure oxygen, in sintering time from 20 to 80 hr, and temperature from 1150" to 1250°C had no effect on the subsequent reduction behavior. The reduction of the samples took place in a vertical tube furnace in which was placed a 1-in. transparent silica tube. The 24-in. heated length

Jan 1, 1964

By L. M. Pidgeon, P. G. Thornhill

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. lengtn of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of ¼ in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10o to 15oC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press Operating at a temperature of 140°C

Jan 1, 1958

By J. H. Wernick, D. Dorsi, K. E. Benson

A LTHOUGH a considerable number of experi--ti mental investigations dealing with the roasting of sulfide minerals have been reported in the past,'"" the behavior of the single roasting particle does not appear to have received the attention it perhaps deserves. This is understandable in that the natural sulfide counterpart of the slab or sheet of metal normally used in high temperature oxidation studies is difficult to obtain and prepare, and is, in most cases, virtually impossible to maintain in one piece at roasting temperatures. The employment, on the other hand, of static aggregations of small particles in masses great enough to permit evaluation of the roasting reactions by means of thermobalances or gas analyses introduces other complications, such as localized overheating, sintering, and variable gas-solid contact. In this investigation a compromise between the above extremes was attempted. Closely sized (e.g., 30 to 40 mesh) particles of natural sulfide minerals were isothermally roasted, out of contact with one another, in an air-swept system. The progress of the roasting reactions was followed by examination in polished section of the treated particles and by X-ray and chemical analysis of their residual sulfide kernels. In this way some conclusions respecting the mechanism of roasting reactions were drawn. Roasting Furnace—The vertical tube furnace used for roasting the sulfide particles had a heated zone 10 1/2 in. long in which four retractable light gage stainless steel trays were suspended. Each tray, and an accompanying baffle, was mounted on a 2 in. iength of 3/8 in. stainless steel tubing. These assemblies were axially suspended on a 30 in. length of in. stainless tubing, by means of which they could be lifted from the hot zone to a water-cooled brass chamber at the top of the furnace. Since each tray unit could slide freely on the 30 in. tube, any desired number of the trays could be retained in the cooling chamber by the manipulation of a horizontal plunger during a momentary withdrawal of the four trays from the hot zone. Although three heating coils, each equipped with variable external resistance, were employed, it was found impossible to avoid a drop of 10O to 15OC between the temperature at the top tray and that of the lower three. Consequently, only the bottom three trays were used for roasting. Temperature readings inside the furnace were made by means of a thermocouple inserted down the tray support tube at tray levels, and a separate thermocouple, inserted adjacent to the middle heating coil, was connected to a Micromax temperature control unit. Experimental Method—The sulfide minerals were obtained in a form as free as possible from gangue and contaminating sulfides, and were crushed and screened to the required particle size range. The tray assembly was kept inside the furnace during the heating-up period and, when a steady temperature had been reached, the trays were withdrawn. Each tray was sprinkled with from 200 to 500 mg of the sulfide grains, and the assembly was lowered into the furnace without delay in order to minimize the time required to achieve the operating temperature. The furnace top was then clamped into position and the thermocouple inserted down the center tube to the required level. It was found that, when this procedure was used, the trays achieved their former temperature within 2 min, after which time the flow of dry air was begun at a rate of 1 liter per min, and timing of the run was started. At the required intervals of time the trays were successively withdrawn from the hot zone and stored in the cooling chamber by means of the retention plunger. On completion of the run the treated particles were removed from the trays to be set aside for examination. The roasted particles were mounted in Lucite on a Buehler press operating at a temperature of 146°C

Jan 1, 1958

By M. R. Geer, J. D. Davis, H. F. Yancey

ALTHOUGH the mechanical cleaning of coal is carried out at plants on the surface, preparation is actually begun at the face in the mine. Here the character of the coal and the amount, character, and distribution of the impurities within the bed may be seen. Blasting methods, dust-allaying procedures, location of mining cut with respect to bands of impurity, and the practice of discarding segregated impurities underground, all have an important bearing on the successful operation of surface cleaning units. The most important single characteristic of both coal and its associated impurities from the standpoint of cleaning is specific gravity. Other characteristics related in one way or another to the preparation of coal are rank, petrographic composition, structure and hardness, and friability. These and still other coal properties are discussed in the first section of this chapter and impurity characteristics in the second. COAL CHARACTERISTICS RELATED TO PREPARATION Rank As a result of work done over a period of several years by committees of the American Society for Testing Materials and the American Standards Association, standard specifications are now available showing the classification of coals according to rank and quality of grade. The term "rank" expresses the degree to which the original coal-forming material has been changed by metamorphism through successive stages from peat to anthracite. Table 1 shows the classification of coals according to rank as devised by the committees of the organizations mentioned. The basic scheme of classification for the higher-rank coals is according to fixed carbon content, calculated to the dry, mineral-matter-free basis. Groups of lower rank are divided

Jan 1, 1950

By B. M. Larsen, T. E. Brower

IN an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats in different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper.2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, ?[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus permits more definite conclusions to be drawn. This excess concentration of oxygen, ?[0], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration ?[O] in liquid steel varies from about 0.009 to 0.035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following : [ ] as in a reasonably steady state-that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour-?[O] returns from

Jan 1, 1946

By R. S. Handy

Does hand-sorting of mill feed pay? The fact that the practice is so general would seem to indicate that there must be good reasons for following it; yet, to my mind, the advantage in many cases is doubtful enough to invite a thorough discussion of the subject, in an effort to determine under what conditions hand-sorting does pay. A typical sorting plant in the Cœur d'Alene district handles about 800 tons of mine-run ore per day. The material is run over a grizzly with about 4-in. opening; the oversize is crushed to about 4-in. size and joins the undersize, and the whole product is washed and screened on trommels with about 11/2-in. openings. The oversize from these trommels is spread on a belt and the waste and first-class ore are sorted out by hand, the residue being further crushed, and passes, together with the trommel undersize, into the mill feed bins. About 50 tons of shipping ore and 150 tons of waste are sorted out per day, leaving about 600 tons of mill feed. To sort out this material requires 20 ore sorters and five bosses and repairmen. The normal cost of hand-sorting is about 16 c. per ton of run-of-mine ore, or 65 c. per ton of material sorted out. I believe that straight crushing and milling of the run-of-mine ore, in this case, is not only cheaper, but is metallurgi-cally more economical and efficient than hand-sorting followed by milling of the residue. At one of the Bunker Hill rock-houses, 1250 tons of run-of-mine ore are reduced to pass 30-mm., or about 11/4-in., opening in six hours, requiring five men. The equipment and the power consumed are not greatly different from that required in a sorting plant of the same capacity, and the normal cost of thus preparing the feed for milling is under 3 c. per ton of mine product. The relative cheapness of this method as against hand-sorting is thus clearly established, the saving being about 13 c. per ton. If there is any virtue in, hand-sorting, then, it must lie in the metallurgical results achieved. Possibly the favor that hand-sorting has found is due to a misconception as to what happens to the clean shipping ore if it is sent to the mill. If a mill makes, say, an 80 per cent. recovery on its feed, the hand-sorting advocate may assume that if the hand-sorted product be sent to the mill, 20 per cent. of the metals contained in it will be lost. This, I believe,

Jan 1, 1920

By R. W. Borio, T. S. Spicer

Inflammability and crossing point apparatus were used to study the influence of increasing amounts oi such inerts as ash, fly-ash, pyrite, limestone and Portland cement with several pulverized bituminous coals. The results indicate that it requires large proportions of inerts to substantially suppress the ease of ignition. The ignition phase of the combustion of solid fuels has been under study periodically since 1900. With the advent of pulverized coal firing some 60 years ago, a technique now dominating the coal utilization field, the ignitibility characteristics have become increasingly important. Much of the early work was directed toward the prevention of coal dust explosions and improved mine safety rather than combustion. Many factors affecting ignitibility have been studied and numerous methods devised in an effort to measure the ignition temperature and these have been reviewed by other authors. A recent survey was made by Finney and Spicerl in 1959, so a discussion of previous work will be omitted. Despite these efforts, the initial stage of combustion, termed ignitibility, is still a little understood phenomenon. In fact, no ASTM method has been adopted for the calibration of this illusive property. However, a D-5 Subcommittee XVI on Ignitibility of Coal and Coke of ASTM (now inactive) defined ignitibility:' "The ignitibility of a fuel is that char-

Jan 1, 1963

By C.R. Taylor, John Chipman

The study of chemical reactions of liquid steel and basic open-hearth slag involves a complex slag system of at least eight important components, and often a number of others. In initiating an experimental program on the equilibrium relationship of slag and metal, it became evident that but little progress could be expected in the interpretation of the complex slags until an understanding had been gained of the simpler liquid systems composed of a limited number of slag constituents. The simplest system that in any way resembles open-hearth slag must contain lime, silica and iron oxide. Attempts to produce in an induction furnace slags containing only these constituents have met with serious limitations because of the solubility of refractories in the slag. Thus in acid-lined furnaces the slag is always saturated with silica and consequently contains between 48 and 63 per cent of this component. In basic-lined furnaces an undesired additional component, magnesia, is introduced to the extent of some 3 to 15 per cent, depending upon the acidity, sometimes with disastrous results to the furnace. Thus in a magnesia-lined furnace, Fetters and Chipman1 found it impossible to prepare slags in which the ratio (CaO + MgO)/SiO2 was less than that corresponding to the metasilicate (approximately 0.8:I). Between this composition and that of the true acid slags lies a field of composition about which very little is known and which cannot be explored by any ordinary means. This circumstance, as well as the desire to study the more basic slags in the absence of magnesia, calls for the development of a method by which slags may be brought into equilibrium with molten metal without coming in contact with refractory. Such a technic has been provided by the rotating induction furnace developed by Barrett, Holbrook and Wood.2 Three modifications of the rotating furnace were described by these authors. The first was essentially a rotating refractory cylinder containing the magnesia crucible, which was itself heated by a graphite crucible. In the second, only the magnesia crucible rotated, the cylindrical graphite heater being stationary. In the third, the coil and crucible revolved, and the metal was heated directly by induction. Experimental Method Furnace The furnace used in this investigation differed from previous designs and may be Sci- a fourth modification of the original rotating furnace. Its essential fea tures are shown in Fig. I' The Outer case A is a fire-clay flue liner for control of the atmosphere. The top B and base 6 are alberene stone. The coil supports D are of ""Formica," a material formed from phe-

Jan 1, 1943

By C. R. Taylor, John Chipman

The study of chemical reactions of liquid steel and basic open-hearth slag involves a complex slag system of at least eight important components, and often a number of others. In initiating an experimental program on the equilibrium relationship of slag and metal, it became evident that but little progress could be expected in the interpretation of the complex slags until an understanding had been gained of the simpler liquid systems composed of a limited number of slag constituents. The simplest system that in any way resembles open-hearth slag must contain lime, silica and iron oxide. Attempts to produce in an induction furnace slags containing only these constituents have met with serious limitations because of the solubility of refractories in the slag. Thus in acid-lined furnaces the slag is always saturated with silica and consequently contains between 48 and 63 per cent of this component. In basic-lined furnaces an undesired additional component, magnesia, is introduced to the extent of some 3 to 15 per cent, depending upon the acidity, sometimes with disastrous results to the furnace. Thus in a magnesia-lined furnace, Fetters and Chipman1 found it impossible to prepare slags in which the ratio (CaO + MgO)/SiO2 was less than that corresponding to the metasilicate (approximately 0.8:I). Between this composition and that of the true acid slags lies a field of composition about which very little is known and which cannot be explored by any ordinary means. This circumstance, as well as the desire to study the more basic slags in the absence of magnesia, calls for the development of a method by which slags may be brought into equilibrium with molten metal without coming in contact with refractory. Such a technic has been provided by the rotating induction furnace developed by Barrett, Holbrook and Wood.2 Three modifications of the rotating furnace were described by these authors. The first was essentially a rotating refractory cylinder containing the magnesia crucible, which was itself heated by a graphite crucible. In the second, only the magnesia crucible rotated, the cylindrical graphite heater being stationary. In the third, the coil and crucible revolved, and the metal was heated directly by induction. Experimental Method Furnace The furnace used in this investigation differed from previous designs and may be Sci- a fourth modification of the original rotating furnace. Its essential fea tures are shown in Fig. I' The Outer case A is a fire-clay flue liner for control of the atmosphere. The top B and base 6 are alberene stone. The coil supports D are of ""Formica," a material formed from phe-

Jan 1, 1943

By G. L. Boxer, C. B. Smith, W. J. Atkinson

Since the inception of the Ashton (formerly Kalumburu) Joint Venture in 1972 the parties have expended $24 million in exploration for diamonds in the Kimberley region. This has resulted in the discovery of the currently subeconomic Ellendale deposits in 1976/77 and the economic Argyle - Smoke Creek Deposits in 1979/80. In addition to Argyle and Ellendale, some 90 kimberlitic and lamproitic bodies have been discovered grouped broadly into three provinces, on the west, east and north margins of the Kimberley block. These bodies range in size from dyke-like features less than 1 m (3 ft) in width to pipes with a surface area of 128 ha (316 acres). The bodies with larger surface area are volcanic crater deposits, infilled with air-fall and water deposited tuffs and epiclastic sediments. Diamond content ranges from trace to economic concentrations. The exploration discoveries were facilitated by early recognition that the petrography, mineralogy and chemistry of the diatremes varied from the classical kimberlites of Kimberley, South Africa, to unusual more highly fractionated, alkaline, silicic varieties having affinities with leucite-lamproite. Indirect methods of exploration were used, primary reconnaissance work being stream gravel sampling for kimberlitic indicator minerals. Evaluation showed Ellendale diamonds are 90% gem and near gem, but the low grades and tonnage render the deposits unpayable at today's diamond prices. After discovery in late 1979 and completion of a feasibility study by the end of 1982, evaluation studies proved Argyle to be economic. Alluvial diamond mining commenced in 1983 producing 6.2 million carats (Mct) that year, 60% being of industrial quality. Open-pit mining of the AK1 lamproite should start in 198511986 at a rate of 3 million t/a (33 million stpy) and produce 25 Mct/a, making this mine, and Australia, the world's largest producer of diamonds by weight. The revenue from diamond deposits is a function of the grade and the value of the diamonds. The grade is the product of the frequency of occurrence of the diamonds (stone density) and the mean stone size; the value of the stones is strongly related to their size distribution. During evaluation of the West Australian deposits, the diamond distribution was found to vary markedly in stone density and stone size. These variations corresponded with changes in the rock type hosting the mineralization. As a consequence, geological mapping of individual lamproites within the multiple vents and careful monitoring of the diamond distribution became an important part of the evaluation work. Similar variations in stone size and frequency were found to accord with separate alluvial terraces and fan gravels derived by erosion from the Argyle AK1 lamproite.

Jan 1, 1985

By W. E. Gaby

SINce the time of their discovery, the genesis of the ores at Mount Morgan, and the nature of the changes which have affected the surrounding rocks, have been the subject of investigation and speculation by geologists and mining engineers. Developments at this mine, for a long period the greatest gold producer of the world, show that with depth the orebody grades into copper, of which metal it now furnishes a large output, exceeding in value the gold. In earlier times, the purity of the gold and the first extremely heterogeneous character of the ore made it an interesting problem. Now, the probable depth of the copper enrichment, genesis, and true relations of the neighboring rocks furnish equally profitable material for investigation, and the present somewhat detailed description of these ores and rocks, from microscopic study, may prove of interest. , The mine occupies a highly silicified portion of a belt of quartz-porphyry, which is believed to be the oldest rock of the region. This belt of quartz-porphyry and allied rocks runs north and south between two areas of a later intrusive granite, as shown on the geologic map,' Fig. 1. Near the orebody the quartz-porphyry has been intruded by a dense basaltic mass called the " Old Basic." Cutting all three of these igneous masses are two sets of basaltic dikes, one running northwest-southeast, and the other, cross dikes, northeast-southwest. These dikes may be contemporaneous with, instead of later than, the ore. Their number has caused peculiar difficulties in extracting it, as the clay selvage of decomposed material affords no adhesion between dike and ore, causing great blocks to loosen and come down. The investigation was based on specimens furnished by B. Magnus, former general manager of the mine, and notes on the areal geology by

Jan 1, 1917

By C. Ernest Birchenall, Robert F. Mehl

Application of diffusion laws in the customary form to experimental studies in binary metallic solid solutions has shown the diffusion cocfficient to vary with concentration for all systems investigated. The factors determining this variation have not been ascertained. However, the question of whether the diffusion coefficient would become a constant, independent of concentration throughout a single phase if the concentration gradient in the diffusion equation were replaced by an activity gradient has been an attractive subject for speculation. Wagner¹ derived an expression for the force acting on a diffusing particle in terms of the chemical potential. This was also the approach chosen by Jost2 in his monograph. From a somewhat different point of view, Eyring3 and his co-workers have used activity gradients when dealing with diffusion in concentrated nonideal solutions. The only actual test with experimental data is that given by Darken,4 though this is more qualitative than quantitative. Many others opposed the application in any way of thermodynamic concepts to kinetic phenomena, including diffusion processes. A rigorous examination of this principle requires good data expressing the relationship between activity and concentration for solid solutions for which reliable diffusion work has been done. The literature provides such information only for the interstitial solution of carbon in gamma iron (austenite) and for the substitutional copper-zinc solutions (alpha brass). Activity data for carbon in austenite at 800°C and 1000°C are available from the recent work of Smith5 on the equilibrium of hydrogen-methane and carbon monoxide-carbon dioxide mixtures with gamma iron. Combining these with the diffusion studies of Wells and Mehl6 makes possible a quantitative test for this interstitial solid solution. The vapor-pressure determinations for zinc over brass by Hargreaves7 allowed the calculation of activities in this system over a wide range of temperatures and concentrations. Extensive investigation of diffusion in this substitutional solid solution has been carried out by Rhines and Mehl.8 Activities in Brass9 If the vapor phase of zinc is assumed to behave as an ideal gas, any solid or liquid solutions that are in equilibrium with the same partial pressure of zinc at the same temperature have the same thermodynamic activity of zinc. In order to assign a numerical value to the activity azn, it is necessary to choose some standard reference state to which an activity value of unity is arbitrarily assigned. It is customary to choose the liquid or solid state of the pure material. Then azn is defined numerically by

Jan 1, 1947

By C. Ernest Birchenall, Robert F. Mehl

Application of diffusion laws in the customary form to experimental studies in binary metallic solid solutions has shown the diffusion cocfficient to vary with concentration for all systems investigated. The factors determining this variation have not been ascertained. However, the question of whether the diffusion coefficient would become a constant, independent of concentration throughout a single phase if the concentration gradient in the diffusion equation were replaced by an activity gradient has been an attractive subject for speculation. Wagner¹ derived an expression for the force acting on a diffusing particle in terms of the chemical potential. This was also the approach chosen by Jost2 in his monograph. From a somewhat different point of view, Eyring3 and his co-workers have used activity gradients when dealing with diffusion in concentrated nonideal solutions. The only actual test with experimental data is that given by Darken,4 though this is more qualitative than quantitative. Many others opposed the application in any way of thermodynamic concepts to kinetic phenomena, including diffusion processes. A rigorous examination of this principle requires good data expressing the relationship between activity and concentration for solid solutions for which reliable diffusion work has been done. The literature provides such information only for the interstitial solution of carbon in gamma iron (austenite) and for the substitutional copper-zinc solutions (alpha brass). Activity data for carbon in austenite at 800°C and 1000°C are available from the recent work of Smith5 on the equilibrium of hydrogen-methane and carbon monoxide-carbon dioxide mixtures with gamma iron. Combining these with the diffusion studies of Wells and Mehl6 makes possible a quantitative test for this interstitial solid solution. The vapor-pressure determinations for zinc over brass by Hargreaves7 allowed the calculation of activities in this system over a wide range of temperatures and concentrations. Extensive investigation of diffusion in this substitutional solid solution has been carried out by Rhines and Mehl.8 Activities in Brass9 If the vapor phase of zinc is assumed to behave as an ideal gas, any solid or liquid solutions that are in equilibrium with the same partial pressure of zinc at the same temperature have the same thermodynamic activity of zinc. In order to assign a numerical value to the activity azn, it is necessary to choose some standard reference state to which an activity value of unity is arbitrarily assigned. It is customary to choose the liquid or solid state of the pure material. Then azn is defined numerically by

Jan 1, 1947

INTRODUCTION The preceding chapters have brought us to a point in the mineral appraisal process where the composited sample values can now be used to estimate the grade and tonnage of the total mineral deposit. There are many techniques whereby drill hole composites and/or other sample data are used to assign grade and estimates to the surrounding rock. Matheron and his students have properly called such techniques extension functions. Before embarking upon a detailed discussion of extension functions, it is necessary that we take time to consider the concept of geostatistics as advocated by George Matheron who developed the "Theory of Regionalized Variables," which he published in 1962 (Matheron, 1971). The concept of regionalized variables has important application in the techniques of mineral appraisal. The position or location of a mineral sample within a deposit can be almost as important as its value. Geostatistics, when defined as "the application of the theory of regionalized variables* to the study of mineralized volumes of rock," takes into consideration the position as well as the magnitude of the value of the sample. The principles of geostatistics, as outlined by Matheron, represent a major step forward in the mineral evaluation process, because they provide a sound theoretical and practical basis for quantifying the geological concepts of (1) area of influence of a sample, (2) the continuity or lack of continuity of mineralization within the ore body, and (3) the lateral changes in mineralization according to the trend direction of an ore body and its orthogonal components; or in other words, a measure of the anisotropy of the deposit. Royle (1971) has defined the objects of geostatistic as being: "1. To estimate the most likely value of blocks of ore, or the values of the whole deposit; and 2. to estimate the errors of such estimates. This latter is important as, in addition to providing a check on unwarranted optimism, it shows where more valuation work may be needed." Because of the great importance of these concepts in mineral evaluation, an introduction into the theory of regionalized variables, and the study of a practical approach to its application is outlined here. REGIONALIZED VARIABLES VS. RANDOM VARIABLES Classical statistics, as discussed in earlier chapters, grew out of the study and application of probability theory. Probability is concerned with the outcome of random events and therefore classical statistics is built upon the theory of random variables, i.e., the probability of independent events. The validity of most statistical inferences depends upon sample randomness being assured for unbiased results. Where independent events cannot be assumed the results of statistical inference becomes suspect. On the other hand, geostatistics is concerned with variables that have position in space as well as magnitude, hence the term regionalized variables. Mineral deposits are not the only field to which the theory of regionalized variables is applicable. Many natural phenomena that demonstrate a relationship between change in magnitude and position in space are amenable to application of geostatistical techniques. Mineral deposits vary greatly in their geosta-

Jan 1, 1980

By Wm. D. Owsley, R. E. Moeller

In the petroleum industry, the process known as oil well cementing is the equivalent of pressure grouting in the mining and construction industries. The science of oil well cementing has been known and practiced for many years. At the present time it has been brought to a high degree of perfection in the handling of many different types of jobs, all of which are known as oil well cementing. In order to understand more easily these processes, we will discuss briefly several different types of oil well cementing jobs, the purposes for which they are performed, and the methods of their accomplishment. It is believed that some of .these procedures, in a modified form, can be applied to pressure grouting projects in general. Originally, oil well cementing consisted of placing a neat cement slurry about a steel casing set in a bore hole. Cement is used about these casings to protect them from the corrosive effects of ground waters, to protect the oil and gas zones from encroachment of undesired water from other levels, to prevent the blowing out of high pressure oil and gas zones from shallow levels, and to strengthen generally and support the casing being placed in the bore hole. After the bore hole is completed to the desired depth, casing is run into the well, pumps are attached to this casing and circulation of the mud fluid in the hole is established; this circulation being from the inside of the casing into the annulus about the casing, and thence back to the surface. Next, a cement slurry is mixed and pumped into the inside of the casing string, A plug, which acts as a piston within the casing, is placed above the cement slurry. This is to separate the top of the cement slurry from mud or water which is used to shove the cement down through the casing and out into the annulus. This plug further serves the purpose of shutting off against a baffle in the bottom end of the casing string, thus stopping the flow of cement out of the casing and leaving the shoe, or lower end of the casing, surrounded by cement and also filling the annulus there above. Casing strings cemented in the above described manner may vary in length from a few feet of large diameter pipe to very long lengths of smaller pipe. As of the present date the longest string of casing which has yet been cemented is some 16,600 ft deep. As a general rule, the deeper the well and the more cement slurry used, the higher the pressures which are required to place the slurry. In parts of West Texas, it is not uncommon to use 4000 sacks to cement a single string of casing in place. Machinery used to mix the cement and place it into the casing is mounted on heavy duty, powerfully equipped trucks which are capable of negotiating any type of highway or country roads. These trucks

Jan 1, 1949