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By George Burrell

J. S. HALDANE, M. D., F. R. S., Oxford, England (communication to the Secretary?).-The valuable paper of Messrs. Burrell and Gauger is of special interest to me as, through the courtesy of the management, I had an opportunity of visiting the Portland mine, Cripple Creek, in 1911, in connection with the investigations on acclimatization to high altitudes, carried out on Pike's Peak, near Cripple Creek, by Professors Yandell, Henderson and Schneider, Dr. Douglas and myself.1 The analyses of one or two samples of the vitiated air were entirely confirmatory of those of the authors. A fatal accident owing to two men being lowered down a shaft filled with the gas had occurred in a neighboring mine just before our visit.

Jan 12, 1916

By Mark R. Lamb

The treatment of slime is of special interest to those engaged in cyaniding gold- and silver-ores. The usual practice is to make as small a percentage of slime as possible. In many instances the slime is given no treatment, but is impounded in dams in the hope that the future will develop some method of economically treating this product. The filter-press was the first step upward from ordinary decantation, but, on account of heavy labor-charge and high cost, its use has been limited to high-grade material. The slime at the Combination mill at Goldfield averages perhaps $20 per ton, and although the values are quickly dis solved the filter-press installation was not entirely satisfactory, resulting in the erection of the canvas-cell filter, developed by Chas. Butters and his staff. This filter is a great improvement over the filter-press, and the following description of it will be of interest to those engaged in treating slime produced in crushing ores, especially in view of the fact that, by the use of this filter, slime can be treated at a lesser cost and with a higher percentage of gold- or silver-extraction than in the ordinary treatment of sand; and, moreover, the initial outlay for an all-sliming plant is less than that for the ordinary sand- and slime-plant. The economy of construction of the Butters filter-plant is clearly shown in Figs. 1 and 2, which illustrate the 20-frame filter of 40 tons capacity, now being installed by the Nevada Goldfield Reduction Co. In this construction the slime-pump is so connected that it call pump to or from either tank or the filter by changing the valve-settinga—a combination which is necessary in this particular installation, since sufficient fall is not available for filling and discharging the filter by gravity. Ordinarily, a slime-plant comprises: a filter-box with frames

Jan 1, 1908

By Dennis A. Clark, Brent Larsson

Tunnelling today is a lot more than drilling and blasting. The developments in drilling and blasting technique have made it pos¬sible to save costs both by increasing the advance per round and optimizing the demands on the final contour. Increasing the hole depth by using longer feeds and bigger dril¬ling rigs make the drilling time shorter compared to the set up time. It is important that the deeper holes are followed up by using a suitable cut. The most common cut in Sweden today is a parallel hole cut with one or more large holes; with a diameter corresponding to the hole depth and wanted advance per round. It is often specified that smooth blasting is to be carried out with a specific drill hole spacing. It is extremely rare, however, for specifications to be issued concerning the alignment of the raw of holes closest to the contour holes. It is very easy to ruin the effect of smooth blasting altogether by firing the rest of the round carelessly with an excessively high charge distribution. The drilling and blasting of the contour holes and the line of holes closest to the contour holes is therefore an extremely important part of any underground work. Accuracy and care in connection with the drilling of contour holes is also extremely important since it is possible to reduce the amount of overbreak, the result being more rapid mucking out, smoother and more durable surfaces and also less consumption of concrete for lining purposes.

Jan 1, 1983

Secondary minerals are the result of a process of concentration and enrichment and are commonly richer than the primary minerals of the same deposit. Secondary ores that contain abundant sulphides are commonly distinguishable in the field from primary ores; this distinction is more difficult with ores that carry a small quantity only of sulphide minerals, and microscopical examination of the ore in thin section is advisable in all doubtful cases. It is frequently difficult, and in some cases impossible, to determine the primary or secondary character of quartzose gold ores. The safest guide is the presence or absence of gaseous or fluid inclusions in association with the ore minerals; such primary association coupled with a lack of traces of oxidation is indicative of primary origin. Any ore that carries traces of oxidation, such as iron stains, endritic oxides of manganese, kaolin and so forth must be considered to have been acted upon by surface agencies, and secondary enrichment must be suspected. Further examination, however, may indicate that the contained value is primary, and, therefore, persistent in its vertical distribution. The manner of occurrence of secondary minerals is frequently characteristic; secondary minerals, being due to the action of surface waters, are commonly found in cracks through earlier minerals, or as crusts lining vugs or surrounding nucleal masses, or in general, in some distribution later than and not conforming to the arrangement of the primary minerals. Secondary sulphides are often found as soft black pulverulent masses or coatings, although they frequently occur well crystallized or as

Jan 1, 1932

By E. C. Truesdale, R. K. Waring

THE Research Division of The New Jersey Zinc Company (of Pa.) has conducted, over a period of years, numerous tests of the reducibility of various zinc ores and the reactivity of various kinds of coal, using a number of different experimental methods. Most of these tests have involved the reduction of mixtures of ore and coal. In such cases the following simultaneous reactions are believed to be mainly responsible for the production of zinc, as was shown by Bodenstein1 some years ago: ZnO + CO = Zn + CO2[I] C + CO2 = 2CO[2]. Each of these reactions is dependent upon the other for the supply of CO and C02, respectively, and the over-all rate of formation of zinc vapor will be determined by the slower of these two reactions. An analogous pair of reactions sometimes of importance is: ZnO + H2 = Zn + H2O[3] C+ H2O=CO+H2[4] While it is possible to obtain an indication of the relative reducibility of various ores, or of a fine versus a coarse ore, by testing them with a given coal, and while coal reactivities may be measured in a corresponding manner, a more direct determination of relative reaction rates is possible by carrying out the desired reactions, such as I and 2, separately and under reproducible conditions. It is one purpose of this paper to present some of the results that have been obtained by the latter method. C. G. Maier2 discusses the relative rates of reactions I and 2 and concludes that while at relatively low temperatures reaction I is faster than reaction 2, this difference becomes less with increasing temperatures, although at smelting temperatures reaction I is still the faster. He also says (p. 48), "At temperatures . . . between 1100° and 1300°C the zinc oxide reduction reaction is intrinsically more rapid than its physical limitations; that is, it is limited by gas diffusion rates." In the course of numerous reactivity tests involving the separate use of reactions and I and 2, we have obtained data that show by direct experiment, we believe for the first time, the correctness of Maier's statements concerning these reactions. We also include data on the relative rate of reaction 3, involving reduction by hydrogen. Finally, some interesting results obtained by carrying out the reduction of a zinc ore-coal briquet in streaming atmospheres of nitrogen, carbon monoxide and hydrogen, respectively, are presented.*

Jan 1, 1941

(Members are urged to send in for this column any notes of interest concerning themselves or their fellow-members.) Members and guests who registered at Institute headquarters during the period Oct. to Nov. 10, 1915: Lester S. Thompson, New York, N. Y. Arthur L. Sweetser, Boston, Mass. William Reed, El Oro, Mexico. Dorsey A. Lyon, Salt Lake City, Utah Heinrich Ries, Ithaca, N. Y. D. W. Blaylock has resigned his position with the St. Joseph Lead Co., Bonne Terre, Mo., and has accepted the position of division engineer for the Big Creek Coal Co., of Chicago, at Harrisburg, Ill. R. B. Lamb announces the removal of his office to Room 1809, 43 Exchange Place, New York, N. Y. C. C. Sanders has taken a bond and lease on the Red Mountain group of 34 mining claims (copper), Patagonia, Ariz., and has commenced extensive development.

Jan 12, 1915

By R. D. Reiswig, J. M. Dickinson

The 0s-Ir equilibrium diagram was determined. The diagram is of the simple peritectic type, with a peritectic temperature of about 2660°C. The solid miscibility gap is narrower than previously reported, the nearly vertical solvus lines occurring at about 42 and 63 wt pct 0s. Widmanstaetten figures resembling martensite in some of the two-phase alloys account for a previously suspected transformation. In view of the long history of use of such alloys, it is surprising that the 0s-Ir equilibrium diagram has not yet been completely determined. The boundaries of the miscibility gap in this system have been estimated to be between 24 and 39 wt pct 0s and 79 and 100 wt pct OS. It has been predicted3 that alloys of ruthenium, osmium, and rhenium with the fcc platinum metals should solidify peritectically in their two-phase regions. The metallographic examination of natural osmiridium alloys has been said4 to indicate, by the presence of a martensitic structure, the existence of a transformation or a phase separation. On the basis of superconducting transition temperature measurements, a 70 at. pct 0s alloy was found5 to be two-phase prior to annealing. EXPERIMENTAL The elemental components used in this work were obtained from several suppliers as sponge. The purities, stated by each supplier to be in excess of 99.8 pct, were confirmed in each case by a semiquantitative spectrographic analysis. The pure components were premelted in an inert-gas tungsten arc furnace with the evolution of some fume, crushed, and pickled for use as alloy melting stock. Upon arc melting to the final compositions, no fume evolution was observed. Because weight losses of the order of 15 mg or less were observed in preparing alloy buttons weighing 15 g, "synthetic analyses" were used in determining the compositions of the alloys. Alloys used in this investigation contained 15, 20, 33, 40, 42, 43, 45, 50, 55, 60, 63, 64, 66, 70, 79, and 80 wt pct 0s. Heat treatments of the alloy buttons and pieces thereof were performed in a previously described6 tungsten tube furnace under pressures of 105 torr or less. It was found that homogenization treatments on the specimens were essential to remove the severe coring which was present, a 3-hr treatment above 2300°C removing all the metallographi-cally visible evidence of coring. Following the various heat treatments, the specimens were "vacuum-quenched" by melting a tungsten fuse wire and allowing them to drop onto a water-cooled copper plate. In a few instances, specimens were quenched into molten tin for comparison, but no difference was apparent. Vacuum-quenched specimens cooled from 2400°C to below a red heat in less than 60 sec. Typical equilibration times varied from 3 hr above 2300°C to 88 hr near 1100°C. The utilization of successively longer equilibration times and the approach to equilibrium from both directions indicated that the times used were sufficient to give a close approach to true equilibrium. Heat-treatment temperatures were measured with a calibrated Leeds and Northrup optical pyrometer through a 0.120-in. lateral sight hole in the tungsten heater tube. It was known that the assumption of black-body conditions inside the 1-in.-diam by 10-in.-long tube was unwarranted. The empirical results of work with thermocouples and melting

Jan 1, 1964

By Oliver C. Ralston, Foster Fraas

INTRODUCTION ELECTROSTATIC methods of separation are used only when some peculiar advantage is gained. Such cases are minerals that are not separable by differences in specific gravity or magnetic response and that are desired in particles as large as 10-mesh, a size on which flotation separation is inefficient. This paper covers several separations of this type and others. It includes studies in the electrostatic behavior of kyanite, sillimanite, and zircon, simple silicates that cannot be called either acid or basic but are more nearly neutral. Kyanite and sillimanite are both desired in 10-mesh size and are difficult to beneficiate from low-grade ores to give that type of concentrate. A separation of garnet from diopside was studied because this mixture had been removed from a wollastonite product by drying and magnetic separation, and no special or extra-preparation for electrostatic separation was necessary. THE SEPARATOR The separator used in these tests is of the nonhomogeneous field type shown in Fig I[a]. It consists of a 11.4-cm-diam rotating metal roll 2 at ground potential over which the dry mixture of minerals is fed. The roll serves as one electrode of the separating field, and the other electrode is that of 3, a smaller 5-cm-diam rotating metal cylinder which is held at a high direct-current potential. In this paper the term "electrode" will be confined to the electrode at 3, roll 2 being designated only as "roll." [ ]

Jan 1, 1947

By A. G. McGregor

IN THE Southwest a number of large steel chimneys discharge the gases from the copper smelting furnaces. Some of these chimneys show no deterioration after twenty years, others show serious deterioration after tour years service. A steel stack 20 ft. 7 ½ in. (6.3 m.) in diameter by 279 ft. (85 m.) high, Fig. 1, used for roaster gases only at the Calumet & Arizona Mining Co.'s plant at Douglas, Ariz., was lined to the top of the bell, 40 ft. (12 m.) from the base, with common brick placed on end, lime and cement mortar being used. Above the bell, the stack was lined with 4 by 4 by 8 in. (10 by 10 by 20 cm.) building tile, placed on end, making a lining 4 in. thick. The tiles were supported on angle-iron rings, riveted to the inside of the chimney shell about 15 ft. apart. The chimney was put into use in the latter part of 1913. Four years later, considerable of the tile lining had fallen out of place and mixed with the flue dust at the bottom of the chimney. About a year later, several . flattened and distorted sheets were noticed about 70 ft. below the top. An investigation showed that the thickness of these distorted sheets, which was originally ¼ in., had been reduced one half. One year later, or six years after the chimney went into use, several holes, a square foot and smaller in area, appeared near the flat spots mentioned. Soon afterwards, when the chimney was shut down for repairs, it was found that most of the tile lining for 125 ft. from the top had fallen out and that the steel shell was badly corroded; but where the lining remained in place the shell was undamaged. Many of the tile which had fallen out of place had become soft or flaky and crumbling, others were as hard as when placed. An analysis of the hard and soft tile is shown on p. 2. These tile were of a fireclay mixture, and were used instead of common brick on account of lower cost. The steel forming the upper 125 ft. (38 m.) was replaced, and a 4 ¼ -in. (11-cm.) lining of firebrick was laid from the bell to the top of the chimney. The brick were placed on end, wedges and keys being used. After each 8-in. course was set, keyed, and wedged tight, thin mortar -water, sodium silicate, and finely ground silica-was worked into the joints. The brick are supported on angle-iron rings, every 15 ft., in the same way as the tile.

Jan 1, 1921

By Yap Chu-Phay

IN 1923, when the writer began his studies in metallurgy, he came upon an article by Professor Honda on the, Equilibrium Diagram of the Iron-carbon System, 1 which made a profound impression on him. What interested him most was this sentence: "In the recent diagram of the iron-carbon system, the solidus is usually shown as a curve bending considerably downward. But it can be shown by thermodynamics that if the liquidus is straight or slightly curved, the same must also be true for the solidus. Hence, in my constitutional diagram, the lower portion of the solidus is slightly bent upward." Ever since then the writer has been interested in thermodynamics, especially chemical thermodynamics. The applications of thermo-dynamics to the metallic systems have been, in the writer's opinion, sadly neglected, although we continually use the phase rule in our study of equilibria in the liquid and solid states of metallic alloys. Many investigators, among whom may be mentioned Heycock and Neville, and s Honda and his coworkers, have applied the simple Raoult-van't Hoff law of depression of freezing point, mainly to determine the heat of fusion, but no attempt has been made to investigate the applicability of the thermo-dynamic laws of solution (as opposed to the laws of solution derived from other sources) to metallic systems" showing either limited or complete solid solutions. Perhaps the reason for this neglect is that until recently there have been so few reliable data on the solidus, which is much more difficult to determine than the liquidus. The writer has under way at the present time an investigation of a number of systems, the solidi of which have been more or less reliably determined. Recently, the writer has not only shown the valid applicability of the thermodynamic laws of solution as applied to the system antimony-bismuth but has also suggested a method of attacking problems involving complete and incomplete equilibria.2 In the present paper, the same method of approach is employed. For convenience, we shall divide this paper into two parts:

Jan 1, 1931

By W. Francis

OF THE methods that have been used for studying the chemical composition of coal, attack by reagents has not, in general, yielded much information. Most of the reagents used have been strong oxidants such as nitric acid or Schulze's solution. The use of a milder oxidizing agent, such as air at low temperatures, enables progressive changes in the character of the coal substance to be studied and thus helps toward an understanding of its constitution. British bituminous coals frequently have a banded structure. These bands (of four types) can be separated one from another and have been found to possess markedly different chemical characteristics, even when taken from contiguous portions of the same lump coal. It has been found convenient to give each of the ingredients of banded coal a distinctive name, more or less descriptive of its appearance, and the study of the chemistry of bituminous "coal" in England has resolved itself into a study of each ingredient separately. . The coal chosen for this work was taken from the Top Hard seam at East Kirkby colliery, Nottinghamshire, a detailed chemical analysis of which has been recorded by Baranov and Francis.1 The banded structure was well marked and the bands could readily be separated. The research was confined to the brilliant vitreous bands (vitrain), the dull hard bands (durain) and the "sooty partings" (fusain); for previous work had shown that the fourth type, to which the name clarain has been given, is intermediate in character, chemically, between vitrain and durain.

Jan 3, 1925

By Dr. O’Neil Thomas J., Donald W. Gentry

Pure gold was hidden in the quartz, they said, 'Twas proved by dreams and signs, and rods divining, By chemic tests, and spirits of the dead, In fact by everything-except by mining. " from R.E. White, "The Mining Town," Alta California, August 18,1873 Textbooks in mining engineering have traditionally covered ore reserve estimation and mine valuation in a single volume. Indeed, Examination and Valuation of Mineral Property by R. D. Parks (1957), having been the standard of the profession for over 20 years, is a classic example. However, the state of knowledge in these two fields has grown tremendously. Most mining engineering curricula now cover these topics in separate courses, and in practice, these tasks are usually performed by different individuals. Thus, this trend now suggests the need for separate, specialized publications rather than a single volume for both topics. Because of the separate development of the science of ore reserve estimation, it is important to state at the beginning what this chapter does not cover. It does not address in detail the design of sampling programs, sampling methods, or the processing of samples or sample data compilation. There are a number of suitable references available on these topics, including Peters (1978), Gy (1979), as well as the previously cited book by Parks. Furthermore, this chapter does not cover in any substantial detail either the traditional methods for using sample data to calculate ore reserves or the more recently developed geostatistical methods. The traditional methods are thoroughly covered by Popoff (1966) and more concisely by Barnes (1980). Among the more useful references for statistical applications in ore reserve estimation are Barnes (1980), David (1977), and Clark (1979). A thoughtful review of the entire science and art of ore reserve estimation is provided by King, McMahon, and Bujtor (1982). In spite of the fact that expansion of knowledge now calls for separate books on ore reserve estimation and financial evaluation, neither publication can ignore the important interrelationships between the two fields. In regard to this text, the unique and crucial role of ore reserves can not be omitted entirely; however, the field is now simply too extensive to treat here in any comprehensive manner.

Jan 1, 1984

By Wittenau, E.

OPERATION of a pilot mill of 100 tons' daily capacity during 1930 and 1931 proved that the copper minerals of the Colorado and Clay sections of the Morenci ore body could be successfully concentrated by the flotation process. Economic conditions caused postponement of further testing work until late in 1938 when the former No. 6 Concentrator at Morenci was converted into an experimental mill of 1500 tons' daily capacity which was known as the Test Concentrator. The Test Concentrator began continuous operation on April 3. 1939, using such crushing and grinding equipment of the No. 6 Concentrator as was appropriate. Modern classifiers and certain flotation machines were installed and later other types of flotation machines, a large grate-type grinding mill. mechanical screens, a short-head cone crusher, and a disk filter were added to the test equipment. Operation was uninterrupted until Dec. 31, 1941.

Jan 1, 1942

By T. Gleason

Problems involved in applying wet scrubbers to gas cleaning coal dust from thermal dryers are reviewed. Careful consideration of all the elements going into a modem coal preparation plant is required in view of the increasing difficulty of applying suitable gas cleaning equipment. This problem arises not only from the presence of finer coals to be handled, but also from the pressures applied by both the general population and the plant operators to improve the standard of gas cleaning performance. The Peabody scrubber is presented by the author as a machine designed to help improve this cleaning operation. Wet scrubbing of fine coal from thermal dryers poses a number of problems in adapting the gas cleaning equipment to the fine particulate matter. Among them are: 1) Dust in gases leaving the thermal dryer cyclones is often extremely fine with high percentages of total dust loading less than 1p in size. 2) Flue gas entering the scrubber is corrosive since the coal fired in the thermal dryer furnace is in the range of 0.7% by weight sulfur or higher. Therefore, comparatively high cost stainless steel alloys or corrosion resistant linings are required for a satisfactory service life. 3) Requirements for residual coal loadings leaving the gas scrubber are continually being lowered, while modern coal preparation operation is imposing heavier burdens on the collection system through introduction into the thermal dryer system of very fine coal previously not recovered. The advent of flotation systems in increasing numbers is typical of this trend. 4) Installation of a scrubbing system on thermal dryers is made primarily to reduce air pollution and product recovery value is of little or no value. Therefore, the operating costs of gas pressure drop and water requirements must be held to a minimum. We have, then, the basic situation which calls for comparatively high cost stainless steel alloy wet scrubbers, which must be highly effective in removing coal fines particularly in the 0 to 5-p size range. Further, the standards of scrubber performance are continually being raised towards the level where gas discharged from the scrubbers must be essentially free of any coloration due to residual dust. In view of the highly refractive nature of fine coal dust, residual loadings of 0.01 grains per cu ft or less are required to provide a color free discharge gas. In the present specifications under consideration a residual coal loading of 0.05 grains per cu ft would be considered color free, since the discharge gases are saturated in the range of 150°F. The steam plume acts as a masking agent at the point of discharge, and upon dissipation of the 150° saturated vapor stream, the effluent gas stream could be essentially color free. The specification under consideration calls for residual solids loading of 0.05 or less grains per cu ft, given a coal size distribution shown in the basic data in Table I. It is noted that the dust laden coal stream has a size distribution by weight of 33% in the 2 to 5-p size range and 55% in the 0 to 2-p size range. The particle size distribution given in the specifications does not detail the percent by weight less than 1-p in size, and more important, the size distri-

Jan 1, 1963

By Charles M. Behre

Under the caption "fault," geologists intend to include all mass movements of solid rocks over adjacent rock masses. When these are studied long after their origin, however, circumstances make it possible to recognize as "fault" only formations showing clear evidence of actual movement of one side with respect to the other. Necessarily the chief evidence consists of the offset of a vein, bed, or other key structure. Planes of movement parallel to the bedding, or nearly so, are recognized with difficulty or are overlooked and hence many important faults are wholly unrecorded. Indeed, among standard American references on structural geology, one alone emphasizes the importance of bedding-plane movements and that one only briefly1. For some time the writer, cognizant of this difficulty, has paid special attention to such movements, both as observed in the field and as recorded in the literature. It is believed that enough data are now at hand to justify a short discussion. Its purposes are fivefold. It is desired first to briefly classify and describe at least the more common kinds of bedding-plane faults, and second, to examine the causes and manner of their formation, at least in instances where the subject has not yet been well covered by other investigators. Third, certain by-products of bedding-plane movements are of such conspicuous importance to the economic geologist and mining engineer as to merit mention. Fourth, concrete illustrations should be placed before the reader in order to lend reality and to furnish a basis of comparison with his experience; such illustrations will necessarily be scattered through the text of this paper. They are only in part from the writer's own field observations. The fifth and prime objective of this paper is to direct attention to the great importance of movements on the bedding planes, in the hope that the recognition of such movements and their significance may simplify mining problems.

Jan 1, 1937

By Frank Firmstone

In the summer of 1842 my father, William Firmstone, was engaged by Charles Jackson, Jr., of Boston, to examine the conditions in the Lehigh valley as a site for blast-furnaces using anthracite for fuel. In consequence of his report, he was further engaged by Mr. Jackson to build a furnace for him and his partners on the Lehigh canal, 2 miles above the mouth of the river at Easton. Work was begun in the fall of 1842, and the first furnace blown-in in 1844. The history of the works, therefore, from 1844 to 1887, when my own connection with them ceased, covers only four years less than the whole period of the rise, culmination, and commencement of the decline in the smelting of iron with anthracite in America, the beginning of which, in a commercial sense, may be put in 1839.' Although this paper, as the title indicates, is concerned almost exclusively with the furnaces at Glendon, still what was done there was more or less influenced by current opinion in the district, and even reflects, to some extent, the changes in opinion and practice with mineral fuel the world over. No. 1 Furnace was built of red brick, on four piers, had three tuyeres and fore-hearth and tymp, as was universal with furnaces using mineral fuel until the introduction of Lurmann's cinder-tuyere. The dimensions and profile are shown in Figs. 1 and 2. The first hearth (Fig. 1) was of sandstone; all after were of fire-brick (Fig. 2). The profile was no doubt derived directly from Gibbons's furnace,2 for it appears from W. Firmstone's note-books that in October, 1840, he visited " Gibbons' new furnaces building at Corbyn Hall," and I have an old drawing,

Jan 1, 1910

By Prakash C. Kapur

Many balling and gradation system6 generate self- similar size distributions of the agglomerates. Inspection of the experimental data shows that this is also true of the steady state distributions resulting from the granulation of coarse, partially wet powders, for example, the fertilizer feeds. This result is exploited in the derivations of a number of simple and useful relationships that are pertinent to the analysis of these agglomeration systems. The bonding liquid demand of the granulating charge is a third order polynomial function of the dimensionless ratio of feed particle size to median granule diameter in the weight-fraction distribution. The median size, in turn, is the scale factor or characteristic size parameter in the similarity distribution. The granule size distribution, therefore, is completely specified by the liquid content and the similarity function, and so are the process efficiency and the ungranulated component In the product at steady state.

Jan 1, 1977

By Frederick L. Knouse

THE south border of Espirito Santo begins about 400 km. north of Rio de Janeiro and extends along the Atlantic Coast northward some 325 km. and inland 100 to 150 km. The area under consideration, where quartz crystal of strategic grades has been mined, extends from 19° to 21° and 30" south latitude; and from 40° to 41° and 30" west longitude. This includes the central part of Espirito Santo, chiefly south of the Rio Doce from the Atlantic ocean, and west across the border into Minas Gerais near the town of Aimores, in the municipalities of Aimores, Minas Gerais; Baixo Guandu, Santa Teresa, Santa Leopoldina, Colatina, Joao Neiva, Fundao and Domingo Martines, Espirito Santo (Fig. 1). During 1943 and the first half of 1944, this region accounted for approximately 3 to 5 per cent of the quartz crystal produced in Brazil. Shipments of crystal have been made chiefly by rail to Rio de Janeiro. This region is serviced by the E. F. Minas-Vitoria and E. F. Leopoldina to Rio de Janeiro through Vitoria; and by the E. F. Minas-Vitoria and E. F. Central do Brasil through Bello Horizonte, Minas Gerais to Rio de Janeiro. The major deposits are accessible from the railheads by dirt roads, some being 5 to 30 km. off the road, reached by trail. In 1943 and the first half of 1944, the bulk of the production was from inter mountain marshes and adjoining hillside alluvial placer deposits in the zone surrounding the town of Santa Teresa. Most of these deposits were formed from the erosion of pegmatites, quartz veins and pockets in the gneiss, with minor quantities produced from miarolitic cavity fillings in the granitic rock. EARLY OPERATIONS Mining of quartz crystal in this region began in 1928, and has been stimulated by the market demand for military uses. Before 1936 little attention was given to it, either in Espirito Santo or eastern Minas Gerais. From 1936 to 1941 the Japanese sent agents to various points of production or potential production to acquire quartz through purchase. They bought crystal as small as 10 grams. Since 1941 only the United States and Great Britain have received quartz from Brazil. The United States purchases crystals of loo grams and larger for strategic purposes. The first known discoveries were made either in the municipalities of Aimores, Minas Gerais, or in Baixo Guandu, or Domingos Martines, Espirito Santo. Marketable quartz crystal was first found either in the massive quartz replacement veins of the Victor Hugo mine in the municipality of Domingo Martines, or in the large cores of quartz in the pegmatites near Baixo Guandu, extending into Minas Gerais. PRODUCTION The production of quartz in Espirito Santo from 1936 through 1941 as recorded

Jan 1, 1946

By Ingvar Janelid

DURING the last ten years, in the effort to save manpower and costs, methods of drilling and blasting in Sweden have changed and developed in a revolutionary manner. These developments have been accompanied by extensive alterations in methods of tunneling and in mining technique. Since the end of the 1940's there has been a 100 pct change-over in Sweden to the use of tungsten carbide tipped drill steel. Both investigations and practical results have shown clearly that a tungsten carbide chisel bit fastened directly on the drill steel is the most economical for Swedish conditions. This applies to drill steel up to the longest size it is practical to handle. In terms of drilling footage, durability of the drill steel and the tungsten carbide bit is generally the same. In this respect, therefore, a detachable bit offers no advantage over integral steel. Transportation of integral steel to the grinding plant, instead of detachable bits only, involves heavier work, but this is counteracted by placing grinding plants closer together or by grinding at the working place itself. The drilling rate with integral steel is higher, the trouble of changing bits is avoided, and the difficulties which always arise with joints are eliminated.

Jan 6, 1954

By L. Guttman, C. S. Barrett, J. S. Bowles

THE transformation from the face-centered cubic (Al) to the face-centered tetragonal (A6) structure in certain alloys of the indium-thallium system reported in the preceding paper1 exhibits many interesting crystallographic and metallo-graphic features, most of which presumably are similar to those of other cubic to tetragonal transformation such as, for example, the one which occurs in the chromium-manganese system,2 the one which, judging by microstructures and X-ray pat-terns,3 probably occurs in the copper-manganese system, the ones accompanying ordering in FePt,4 Copt,&apos; and AuCu,5 and the transformation near 115°C in barium titanate.6,7,8 If, as mentioned in ref. 1, the transformation is second order, the phases in equilibrium do not differ in composition, and it is both possible and likely that the mechanism is a diffusionless one. The present paper reports the results of a detailed study of the indium-thallium transformation and presents a theory for the atomic movements of the transformation which accounts for the observations. Experimental Crystallography of Transformation Markings The transformation from cubic to tetragonal in the indium-thallium alloys produces microstruc-tures of the type illustrated in fig. 1. These structures are observable either as relief effects produced on a smooth surface by the transformation, or after polishing and etching. These microstruc-tures are markedly similar to those found in copper-manganese alloys: iron-platinum alloys," and barium titanate,7,8 where it has been demonstrated that the lamellae are parallel to the (101) planes of the original cubic crystal. That the observed lamellae in indium-thallium alloys are also parallel to the (101) planes was proved by the following analyses performed on single grains in polycrystal-line samples containing 20.75 atomic pct T1. A study of the microstructures of a number of specimens revealed the fact that in many specimens the transformation markings continued unchanged in direction across the (111) interfaces of annealing twins. This at once demonstrates that the lamellae are parallel to a plane that is common to both parent and twin orientations, i.e., a plane in the zone <112>. This conclusion is compatible with the lamellae being parallel to (101) planes. To make a more complete determination of the habit plane of the lamellae, two grains were selected in which annealing twins ({Ill) twins) had been present in the high-temperature modification (cubic). The traces of the annealing twins, and of the transformation markings in both parent and twin crystals, were plotted together in stereographic projection, and a parent-twin pair of orientations was found that accounted for all lamellae as (101) planes. The projection of one of these twinned grains is reproduced in fig. 2. In view of these results the probability of there being another solution for the habit plane remote from (101) is very small, but to explore the possibility that the habit plane deviates slightly from (101), a more precise analysis was undertaken. In this, the orientation of a grain in the cubic modification was determined from a Laue back-reflection photograph taken with the sample at 90°C. The traces of the lamellae that formed in this grain on cooling were plotted in the stereographic projection

Jan 1, 1951