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By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By B. H. Bergstrom, Will Mitchell, T. G. Kirkland, C. L. Sollenberger

In a previous article' the authors outlined a study I of the variables in rod milling and also reported data from a series of open circuit grinding tests on a massive limestone in a 30-in. x 4-ft end peripheral discharge rod mill. As a second part of the experimental program, an analysis is now presented for the 30-in. x 4-ft overflow rod mill grinding under identical conditions, except that discharge ports on the periphery of the mill shell have been sealed so that the products from the present series overflowed through a 9-in. diam opening in the center of the end plate. A variance analysis has been made of the combined data for the two experiments, and performances of the two mills are compared here. Included in the first report were descriptions of feed preparation, rod mill circuit, instrumentation and controls, and techniques used to evaluate data. Dependent and independent variables were defined, and variance analyses were made to test the relative significance of variables and to establish magnitude of error for the experiment. Significant data were plotted in various combinations, and conclusions were drawn from the graphs. The procedure and analysis in this series of tests follows the first tests and is not repeated.

Oct 1, 1955

By Charles R. Keyes

Discussion of the paper of Charles R. Keyes, Bi-Monthly Bulletin, No. 19, January, 1908, pp. 1 to 31. BERNARD MACDONALD, Guanajuato, Mexico (communication to the Secretary*) :-Mr. Keyes's paper is very interesting to me because of my personal experience with the development of the ore-deposits of Lake Valley. Seeing that many of the important facts connected with the early history of the development of these mines are either now unknown, or have become more or less distorted, and thinking that it would be interesting as well as valuable to have these recorded as a historical supplement to Mr. Keyes's elaborate paper, I present the following narrative: In February, 1881, soon after the Atchison, Topeka & Santa Fé railroad made connection with the Southern Pacific at Denting, N. M., the Lake Valley properties were brought to the attention of New York and Philadelphia capitalists by John A. Miller, of Silver City, N. M. At that time silver was $1.10 all ounce and silver-mining was very attractive. Vast fortunes were being made in Leadville and other Colorado camps, and Tombstone was pouring out its millions. Lake Valley, like all' other mining-camps in their early history, had its drawbacks, but the railroad, then passing within 14 miles of it, mitigated many of these. The camp lay in the foot-hills of the Black range, in which the Apaches had their stronghold, and raided, when it suited their convenience, the surrounding country, killing the prospectors and pioneer settlers and stealing their cattle. This band of Indians was joined by the outlaws and desperadoes from the frontiers of Texas, Arizona, and Mexico, who collectively defied the United States soldiers, restrained, as the latter were, by the mollycoddles from the East. The brutality of the Indian raids through Arizona and New Mexico in those days would scarcely be believed now, but they stain in deep

Feb 1, 1909

By John A. Emery, Frank G. Snyder

Mining geology has a threefold objective: to guide prospecting for new ore, to evaluate known orebodies as development risks, and to supply the detailed knowledge of ore structures necessary for more economical extraction. The role of geology in exploration is well established. This article will deal solely with orebody evaluation and the application of geology to development and mining problems as practiced by St. Joseph Lead Co. in the southeast Missouri district, which for many years has supplied approximately 30 pct of U. S. lead production. The primary concern of the present discussion is geology in the district, not geology of the district. The southeast Missouri district, locally called the Lead Belt, includes parts of St. Francois, Madison, and Washington counties. The first recorded mining was about 1720 at Mine La Motte in Madison County on surface showings of lead ore. Later discoveries near Bonne Terre, twenty-five miles to the north, led to extensive prospecting and underground mining. To date the district has yielded well over 8 million tons of pig lead, valued at more than $1 billion. By far the greater part of the lead production has come from an area of 50 sq miles, centering Fig. 1-Central portion of southeast Missouri on the city of Flat River, as shown in Fig. 1.

Dec 1, 1956

By A. J. May

THE Star shaft is near the north boundary of the group of mining claims belonging to the Ground Hog Unit of the American Smelting and Refining Co., near Vanadium, N. Mex. The shaft bins and surface plant are served by a siding from the Atchison, Topeka and Santa Fe railroad, and this factor of railroad transportation influenced the choice of site. This paper describes the shaft sinking only, and does not cover the surface plant or other parts of the mine or its equipment. Experience with this shaft seems of interest because of the mechanical mucking, good rate of progress, and the costs obtained at the present level of labor and material costs. The shaft has four compartments, as shown in fig. 1. Equipment: Riddell Mucker: The mucking was done by a clamshell bucket and a Riddell mucker, which, as in other shafts sunk with this machine, consisted of a carriage mounted on axles with double flanged wheels that traveled back and forth on a swinging track, parallel to the long axis of the shaft (fig. 2 to 6). Three air motors were mounted on the carriage; one hoisted the clam, one operated the bucket opening and closing device, and the third moved the carriage along the track. The carriage was designed with guide shoes and could be hoisted to the surface in either skip compartment without trouble (fig. 4). During blasting operations the clam was tied on the rails and the carriage was hoisted up a few sets to protect it from flying rock. No attempt was made to hoist the carriage and clam together lest the clam catch in the timbers. The track on which the carriage operated was built of extra heavy 6 in. pipe with 1x2 in. strap welded on top for rails. A 1/2x4 in. angle was welded on the bottom of the 6 in. pipe with a 1 in. round rod welded in the angle for a filler and additional strength. The end or cross pieces of the track which kept the track to gauge, were also made of reinforced, heavy 6 in. pipe and were bolted to the long members on which the carriage traveled (fig. 2 and 3). This section of track was used, instead of heavy railroad rails, for two reasons: (1) this section would bend before it snapped and was much safer for the men in the bottom; and (2) it was a better blasting shield, and protected the timbers better, in fact, only five wall plates were broken by blasting in 1926 ft of shaft, although 150 to 225 lb of 40 pct powder was loaded per round, depending on the ground conditions. The swinging track was supported by four 5/s in. hoisting cables anchored to the end plates. There were also four safety cables fastened two to three sets above the carriage, which would catch the carriage if one of the supporting cables should fail for any reason (fig. 7). Clamshell Buckets: Two clams were used, each of 3/s yard capacity, both manufactured by Blaw Knox, Nos. 662 and 663. The No. 663 clam was 4 in. wider and 5 in. shorter than the No. 662, but seemed to dig and load more satisfactorily. It was easier to hoist and lower when it came up for repair. It did not turn over as easily, had less spill

Jan 1, 1950

Reports on process control computer projects at the Kamioka Plant are summarized from the Japanese technical literature (see references). Fluidized Bed Roaster and Sulfuric Acid plant1 A small process control computer (Yokogawa CCS-200 with an OKITAC 5090P 8K central processor) was installed in the Kamioka smelter in the summer of 1964. A year-and-one-half installation project followed, which successfully brought the roaster and acid converters under computer control. The process plant consisted of two 130 RPD fluosolids roasters and two Monsan to acid units. The roasters are charged with lead-zinc concentrates with SO2 from the roaster processed to sulfuric acid. The computer system provides the following functions: 1) log data from the plant--the input is taken from 50 analog signals, 130 high-low digital alarm points, and 50 analog high- low alarm points; 2) set-point control on 18 control loops; 3) direct digital control (DDC) on eight control loops for which conventional controllers were difficult to adapt due to abnormal loop characteristics or delay times; 4) execution of a safe shut-down procedure

Jan 1, 1969

By Molly Gleiser, John Chipman

The standard free energy of formation of WC was obtained from determination of the equilibrium WC + CO2 = W + 2CO between 1215° and 1266°K. Its uallie is -8340 * 300 cal Per mole over the above range of temperature. AS a part of a program devoted to measuring the stabilities of the carbides, a study was made of the free energy of formation of tungsten carbide, WC. The homogeneity range of this compound is very small,' and it is generally accepted and is dealt with here as a stoichiometric compound. The heat of formation, ?H°f, of WC has been accurately measured by Huff, Squitieri, and snyder.2 The only attempt, however, to measure the free energy of formation of this compound was made by Schenck, Kurzen, and wesselkock3 who brought tungsten carbides into equilibrium with H2 and CH4. Unfortunately their results are not sufficiently accurate nor consistent enough to permit the calculation of free energies. METHOD Heating mixtures of tungsten and carbon powders in hydrogen for 4 days at 1050°C led to formation of some WC without any trace of W2C. It was deduced, therefore, that at least below this temperature W2C will not be formed during experiments with mixtures of tungsten and WC. This was subsequently shown to be correct. The method chosen was the measurement of the equilibrium of the reaction

Jan 1, 1962

By Z. T. Bieniawski

Tunneling in coal mines includes driving the main haulageways or roadways which serve as the development entries. Although constituting the lifelines of these mines, little attention has been paid to development entries in the United States in terms of systematic pre-construction planning and design. Coal mine tunnels, as the terminology is used in Europe and elsewhere, are understood in the United States to be the main development entries, either in room-and-pillar mining or in longwall mining. These development entries are not tunnels in the traditional sense since they usually are of a multi-entry type involving entries and cross-cuts, and their intersections. Thus, the design of coal mine tunnels is intertwined with the design of coal pillars. This paper summarizes the results of a three-year research project aimed at improving the design of development entries in U.S. coal mining. A design approach has been developed for this purpose and design procedures for roof support in main entries and for pillar sizing are proposed.

Jan 1, 1984

By R. Floyd Farris

A method is presented for determining minimum waiting-on-cement time, which takes into account the differences that exist between types and brands of cements and such individual well conditions as depth, temperature, and pressure. The basis for the method was determined by laboratory tests. Being a laboratory development, several steps were required to prove its merit. The first step consisted of laboratory tests designed to determine the minimum cement strength required in wells. Basis was found for setting a minimum value of 8 Ib. per sq. in. tensile strength. Next, it was shown by laboratory tests that the time to 8 Ib. per sq. in. tensile strength may be expressed as a function of consistometer stirring time to loo "poises," the approximate relation being "the time to 8 Ib. per sq. in. tensile strength equals the time to I00 'poises' times three." Next, it was shown that the time of maximum temperature development in cement slurries, due to heat of hydration, is also related to consistometer stirring time to loo poises, but only by a factor of approximately two. It was shown also that the shut-in casing pressure will build up after cement is placed and register a maximum pressure at approximately the same time the slurry down the hole attains maximum temperature. From this and the relationships listed above, the general rule was established that minimum waiting-on-cement time (time to 8 Ib. per sq. in.) after casing cement jobs in any well is equal to the time when the shut-in casing pressure reaches a maximum, as measured from the initial mixing of cement, times a factor of 1.5. Cement plugs drilled in the field at the time prescribed by this formula were found to drill "firm to hard," thus confirming the laboratory tests. These tests prove that many of the present regulations for waiting on cement require a longer time than is absolutely necessary. Use of the method herein proposed offers the possibility of a saving of $1200 per well. Introduction The length of time allowed for cement to set after casing is determined either by state-wide rules, field rules, or self-imposed rules written into drilling contracts. In general, the time is dictated by experience and common practice. However, owing to differences in opinion and in experience of the various groups involved, waiting-on-cement time often varies from one area to the next. For example, an operator in an area where no rules exist may drill out of surface pipe at 24 to 36 hr., while another operator in another area may wait 48 hr. or more to comply with state or field rules, although the depth of the well, hole size, type of cement, and other data are identical. An even greater difference in practices will be found by making similar comparisons with respect to oil-string cement jobs. Differences in waiting-on-cement times of 36 to 48 hr. are common. Further complicating the picture is the rather common practice of allowing more waiting time for cement to set at the greater depths than is allowed at the shallow depths. This practice has existed for years in spite of the common knowl-edge1,2,3 that the temperature of the earth at the usual setting depths of surface

Jan 1, 1946

By R. Floyd Farris

A method is presented for determining minimum waiting-on-cement time, which takes into account the differences that exist between types and brands of cements and such individual well conditions as depth, temperature, and pressure. The basis for the method was determined by laboratory tests. Being a laboratory development, several steps were required to prove its merit. The first step consisted of laboratory tests designed to determine the minimum cement strength required in wells. Basis was found for setting a minimum value of 8 Ib. per sq. in. tensile strength. Next, it was shown by laboratory tests that the time to 8 Ib. per sq. in. tensile strength may be expressed as a function of consistometer stirring time to loo "poises," the approximate relation being "the time to 8 Ib. per sq. in. tensile strength equals the time to I00 'poises' times three." Next, it was shown that the time of maximum temperature development in cement slurries, due to heat of hydration, is also related to consistometer stirring time to loo poises, but only by a factor of approximately two. It was shown also that the shut-in casing pressure will build up after cement is placed and register a maximum pressure at approximately the same time the slurry down the hole attains maximum temperature. From this and the relationships listed above, the general rule was established that minimum waiting-on-cement time (time to 8 Ib. per sq. in.) after casing cement jobs in any well is equal to the time when the shut-in casing pressure reaches a maximum, as measured from the initial mixing of cement, times a factor of 1.5. Cement plugs drilled in the field at the time prescribed by this formula were found to drill "firm to hard," thus confirming the laboratory tests. These tests prove that many of the present regulations for waiting on cement require a longer time than is absolutely necessary. Use of the method herein proposed offers the possibility of a saving of $1200 per well. Introduction The length of time allowed for cement to set after casing is determined either by state-wide rules, field rules, or self-imposed rules written into drilling contracts. In general, the time is dictated by experience and common practice. However, owing to differences in opinion and in experience of the various groups involved, waiting-on-cement time often varies from one area to the next. For example, an operator in an area where no rules exist may drill out of surface pipe at 24 to 36 hr., while another operator in another area may wait 48 hr. or more to comply with state or field rules, although the depth of the well, hole size, type of cement, and other data are identical. An even greater difference in practices will be found by making similar comparisons with respect to oil-string cement jobs. Differences in waiting-on-cement times of 36 to 48 hr. are common. Further complicating the picture is the rather common practice of allowing more waiting time for cement to set at the greater depths than is allowed at the shallow depths. This practice has existed for years in spite of the common knowl-edge1,2,3 that the temperature of the earth at the usual setting depths of surface

Jan 1, 1946

By E. I. Organick, B. H. Golding

A simple correlation is presented for the prediction of saturation pressures in condensate-gas and volatile-oil mixtures. Saturation pressure is related directly to the composition of the mixture with the aid of two generalized composition characteristics, B, the molal average boiling point, and W, the modified weight average equivalent molecular weight. The need for hydrocarbon equilibrium constants is entirely eliminated. Only the values of the two conlposition parameters need be obtained, and these are calculated in a simple and reproducible manner from the stream analysis of the mixture and the ASTM distillation for the heavy ends. The correlation is given .in the form of a set of 14 working charts in which saturation pressures, either retrograde dew points or bubble points, are plotted cs temperature, forming partial phase envelopes for mixtures having discrete values of the composition parameters, B and Wul. A locus of critical states on each chart intersects the partial phase envelopes, distinguishing retrograde dew point pressures from bubble pressures. A comparison Between experimental and predicted saturation pressures reveals that: the probable error of the predicted values, when considering the data as a whole (214 experimental points), is five per cent; for high gasjoil ratio condensate fluids (approximately 40.000 or more cu ft/bbl) the probable error is about 8 per cent. In either case, however, the probable error is comparable with the overall uncertainty of the data from which the correlation is derived. INTRODUCTION There is a present need for a simplified, yet reasonably accurate method for predicting upper phase boundaries of reservoir fluids. In the case of high gas/oil ratio fluids, namely the so-called highly volatile oils and the condensate gases, no simple method, accurate within 5 to 10 per cent, is now available by which bubble point and retrograde dew point pressures may be predicted. In order to calculate the upper phase boundary of a volatile reservoir fluid, using correlations now available, it is necessary to find and employ the proper K-values (equilibrium volatility ratios of the hydrocarbons in that fluid). This process of finding the Proper K-values usually involves a double trial and error computation. Since the K-values are in themselves dependent upon the composition of the fluid, their use in the calculation of a saturation pressure actually represents an unnecessary complication. The correlation described in this paper relates saturation pressure directly to the stream composition, entirely eliminating the need for K-values. The only data required to calculate the Stream composition are obtained from the customary low temperature fractional analyses of separator gas and liquidt and from the ASTM distillation of the c,+ fraction of the separator liquid. A considerable simplification in the procedure for predicting saturation pressures is thereby effected without significant loss in accuracy. BASIS FOR A CORRELATION Empirical relationships between the saturation pressure and the composition of the fluid in question have already been partially explored by one of the authors.V n particular, a relationship between the critical pressure for a hydrocarbon mixture and its composition was developed. The critical pressure is in reality a saturation pressure, as it repregents a limiting case either of a dew point or of a buhhle point pressure for any given mixture. It was found that the critical pressure may be correlated as a function of two generalized composition characteristics or variables. One characteristic expressed a molal average prop-erty, the other a weight average property of the mixture. Specifically, the molal average boiling point (MABP, or B as it

Jan 1, 1952

By J. P. Witherow

The plant of Messrs. Oliver Brothers and Phillips, the only one in operation until January, 1886, has not been available for any further experiments since those of Mr. R. W. Hunt, described in his paper on the subject, at the Chattanooga meeting. Continuous working has, however, led to important improvements in plant, and furnished valuable results in practice. The difference between the converter and plant of 1884, as illustrated in our Transactions, vol. xiii., p. 759 and following pages, and those of 1885, may be observed in Plates 1. and 11. accompanying this paper. Plate I. gives the elevation and plan, in half-section, of one of our latest 3-ton converters. Plate 11. gives a plan of the general arrangement of a plant adapted to turn out slabs, blooms, and billets for the general market. The letters on Plate I. refer to the following parts :

Jan 1, 1886

By Richard R. Klimpel

The role of dispersion is described in the wet grinding of ores using tumbling media mills and the flotation of coal, sulfide, and non-sulfide ores. Emphasis is placed on characterizing the effects of dispersion in laboratory tests using engineering models which provide information useful for the improvement and control of existing plant operations. Significant economic gains in the above named operations have been realized using these models as guides for full- scale plant testing.

Jan 1, 1980

By C. W. Allen, B. D. Cullity

The proportional limit of iron crystals in torsion is governed by the resolved shear stress in the most highly stressed slip systems, averaged around the specimen circumference, and does not obey a critical resolved shear stress law. Crystals of most orientations exhibit a stage of easy plastic deformation, akin to easy glide in tensile or shear specimens. Transient deformation, similar to that which occurs in single crystals of other materials, is also observed. THE torsional deformation of single crystals of magnesium (hcp) and aluminum (fcc) has been described recently by Choi et al.,&apos; especially with respect to the criterion for the orientation dependence of the onset of plastic flow in these materials. The purpose of this paper is to present results of torsion tests of iron single crystals and thus to extend this yield criterion to a bcc metal. In addition to considering the variation of proportional limit with crystal orientation, this paper also briefly treats work hardening, transient deformation, and the mechanism of plastic flow in iron. The effects of the method of surface polishing and the chemical purity of the iron have been investigated. STRESS DISTRIBUTION It is convenient to express the stress at any point of a cylindrical crystal stressed in torsion in terms of t0, which is the shear stress acting at the surface on a plane normal to the axis of the cylinder and in a direction tangential to the cylinder. This stress is given by To = 2T/pr3 [1] where T is the applied torque and r the specimen radius. The shear stress t, resolved in any chosen slip system is given in terms of 7, by1 Ts/TO = sin 0, cos d sin (0, -) + cos , sin d sin d - ) [2] where 0 and d are the angles between the specimen axis and the slip plane normal and slip direction, respectively; h is the angular circumferential position on the specimen at which t, is being determined, measured from an arbitrary reference plane which includes the axis;o and d are the angular coordinates of the projections of the slip plane normal and slip direction on a transverse section with respect to this same reference plane. Slip in iron occurs in a <1ll> direction on the {ll0}, (1121, and (123) planes, which together comprise 48 slip systems. A complete evaluation of the stress distribution in an iron crystal stressed in torsion would therefore require a calculation of Ts/T0 as a function of for 48 different slip systems. Fortunately Gough,&apos;who studied the behavior of iron crystals in alternating torsion, was able to simplify this problem considerably. He showed that it was sufficient to consider a kind of average slip plane for each slip direction, namely the mathematical plane of maximum resolved shear stress containing the slip direction considered. This simplifying approximation is possible because, for each slip direction, the active slip plane or planes lie very near this mathematical plane of maximum shear stress. Vogel and rick&apos; have critically reviewed the early work of Taylor and Elam,13 Taylor,14 and Fahrenhorst and schmid8 from which the identification of the above crystallographic planes as slip planes in the bcc lattice largely stems. While their criticism is clearly justified, their own results do little to clarify the issue. The role of cross slip (screw dislocations changing glide planes) is evidently so important in this case, as Read3 has suggested, that methods for deducing slip systems from observations of gross slip traces are inadequate, such traces commonly arising from complex dislocation motion. Thus the treatment given here involving the plane of maximum resolved shear stress seems a logical simplification especially in view of Gough&apos;s2 study of a iron. There is, however, an assumption built into the subsequent treatment the comparative validity of which is difficult to assess, namely, that slip in all slip systems in iron may be characterized by a common critical resolved shear stress. The shear stress 7, resolved in a slip direction defined by d andd, and on the plane of maximum shear stress containing this direction, is found by first maximizing 7s/70 with respect to either Oo or 4,. The slip plane coordinates are then eliminated by using the relation between 0, ,o and d, d, namely,

Jan 1, 1963

By Harold Bloom, H. E. Hawkes

Streams and rivers are the principal channels into which the weathering products of rocks and their contained ores are funneled. The inorganic load of a stream system is a crude sample of all the earth materials in the drainage basin of the stream, and the heavy mineral assemblage of sediments is one of the oldest and most successful guides to certain kinds of bedrock ore in unexplored terrain. Early application of this method was made possible by the ease of separation and identification of heavy minerals in the field with the simplest equipment. Water analysis for traces of base metals became a practical possibility about eight years ago with the development of rapid chemical tests that could be carried out at the field site. Since that time the water method has been widely used in mineral reconnaissance, and a substantial number of discoveries have been reported. Here again, the techniques developed for this work were relatively simple and could be performed easily at the field site.

Nov 1, 1956

By E. Marburg

Most ingots complete solidification vertically rather than transversely. This conclusion is based on complete solidification patterns of big-end-up and big-end-down ingots developed from dumped and split ingots. New mechanisms for segregation in ingots are proposed. Internal soundness in ingots is found to depend upon adequate width of vertical core. A LTHOUGH it has long been recognized that the A internal soundness of ingots is principally been recognizedthatthedependent on the process of solidification, knowledge of this process is still only fragmentary. The mechanisms causing the several types of segregation in ingots, which mechanisms are necessarily related to the solidification process, are not well understood. In absence of a better conception of the solidification process, it seems hardly surprising that mold design varies widely. Guiding principles for mold design are needed. At stake in the improvement of ingot quality are not only improved properties of steel but also appreciably higher yields of sound steel, which would considerably decrease manufacturing costs. In the present study, following a general description of the several types of segregation in ingots, Fig. 1, complete solidification patterns of a big-endup and a big-end-down ingot, including the zones of accelerated solidification, have been reconstructed. In both ingots, a core of vertical solidification extends from the base cone to the hot-top junction. Based on the finding that ingots complete solidification vertically rather than transversely, as has been previously assumed, new mechanisms for segregation in ingots are proposed. Internal soundness in ingots is found to depend upon the width of the vertical core, which in turn depends principally upon the width to height (w/h) ratio of the ingot. Solidification patterns of ingots are of three general types: 1—those in which vertical solidification does not reach the top, 2—those in which vertical solidification just reaches the top of the ingot, and 3—those in which vertical solidification reaches the top some time before transverse solidification would reach the middle of the ingot, so that a vertical core of definite width at the top of the ingot is developed. Axial defects vary characteristically with the solidification pattern. Examples of ingots of each type are presented. Dumped ingot data reported by Nelson&apos; have been used in developing families of curves of transverse and vertical solidification, and in deriving the effect of the w/h ratio upon the solidification pattern. Photographs of many ingots split at Homestead and at Duquesne Works, United States Steel CO., as well as those published by the British Iron and Steel Institute,&apos; have been studied and analyzed. In the interest of brevity, only typical ingots of each type or classification from the above sources have been reproduced. Segregation in Ingots Segregation in ingots falls into three principal classifications: 1—lamellar segregation, as revealed in sulphur prints (Figs. 2 and 3); 2—zone segregation, or broad variations in composition; and 3— interdendritic, or microsegregation. The last class, of minor importance, is not considered here. When liquid metal cools, according to the theory of differential or selective solidification,2 metal of relatively high purity selectively solidifies first. Liquid segregates (principally carbon, phosphorus,

Jan 1, 1954

By S. H. Ash

Safety records in the mineral industry for 1954 will do well to hold their own as compared with 1953, because of the poorer rate in the coalmining branches, even without the recent mine explosion at the No. 9 mine of the Jamison Coal & Coke Co., Farmington, W. Va., that killed 16 persons. .The cause of that disaster has not yet been ascertained. The metal-mining branch appears to have improved its record over the last year, however. Health Diesel equipment continues to be favored in non-coal mines. Equipment includes locomotives, trucks, tractors, and muckers. Of major concern and interest is the establishment of a new threshold limit for oxides of nitrogen-5 ppm as compared with 25 ppm. This gas is the most harmful component of diesel-engine exhaust gas

Jan 3, 1955

By Shiou-Chuan Sun

Most previous work on sedimentation of coal1 and mineral2,3 suspensions has been conducted in graduated 1-liter glass cylinders of 6-cm diam. With this type of large container it is often difficult to see through the suspension and tedious to perform a large number of tests. Furthermore, the heavy glass walls and the graduation lines of glass cylinders cause internal and stray reflections, respectively. To obviate these difficulties, an Atlab Emulsion Viewer from Arthur H. Thomas Co. was adopted as an alternative apparatus. Equipment: The viewer,4 shown in Fig. 1, consists of: 1) a stainless steel frame rack holding 21 flat-bottom Nessler tubes or an ungraduated replica 1.64 cm ID and 1.80 cm OD, 2) a movable masked 150-w lumaline lamp behind the tubes, and 3) a scaled glass panel with 100 vertical divisions in front of the tubes. The tubes are snugly held by springs in slots formed by spacer bars, which also serve as shields to eliminate stray light. The lumaline lamp is mounted behind a slit in a movable housing so that it can be positioned behind the suspension to be examined. The lowest or 100-division line on the front glass plate is leveled with the upper side of the bottom of all tubes. The scaled portion of each tube, roughly 19 cm high and 40 ml in volume, is used to study coal and mineral suspensions.

Jan 1, 1959

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

INTRODUCTION AND SUMMARY BY "thermochemical factors" we refer to those variables which affect the net heat which must be put into the bath in order to make a heat of steel from any given set of charge materials. This is admittedly only one phase of the production rate problem. The whole problem includes also consideration of rate of charging, flame radiation intensity, rate of optimum fuel input obtainable without undue damage to the furnace structure and any variations in conditions such as stirring, slag volume or viscosity which affect the rate at which heat can flow down and away from those exposed surfaces where the heat is first received from the flame. Because a given type of charge in a shop is often considered as being nearly constant in both physical and chemical characteristics from one heat to another, and because such a state of affairs may be thought of as resulting in a reasonably constant net heat requirement, this requirement is usually neglected as an invariable, inevitable matter, about which nothing can be done. With normally variable charge, and variable percentages of limestone and cold pig, however, the net heat requirement can be subject to very appreciable change; it may also be affected by such things as the currently popular procedure of blowing oxygen into the bath, To bring the general problem of production rate into better perspective, therefore, it seems worthwhile to try to evaluate the net bath heat requirement and the amount by which it may change through variation in different factors. Because our calculated values of net heat are admittedly uncertain, we have tried to check them in part by statistical data on actual production rates over rather large groups of commercial heats. The results of calculation of net heat requirements in a wide range of actual commercial practices may first be summarized briefly. We find a wide range between the extremes of different practices, as would be expected, with duplex practice (essentially blown metal plus hot metal) having essentially a zero net heat requirement, an all-cold charge practice somewhat over 1 million Btu per ton of steel and a high hot metal practice about 0.75 million Btu per ton. Over the commercial range of scrap plus hot metal practice, increasing the percentage of hot metal from 40 to 78 pct, with other factors nearly constant, decreased the net heat by only some 5-10 pct, The net heat is increased by about 1.2 pct for each I pct of the metallic charge present as cold pig. Calculated values of oxygen supplied by air varied from about 25 up to 60 lb per ton of steel, and in a 60 pct hot metal practice, an increase from 32 to 52 lb oxygen from air decreased the calculated net heat from 0.88 to 0.74 million Btu per ton. Use of pure cold oxygen to substitute for about one half of the total charge ore (about 28 lb oxygen replacing 145 lb ore per ton of steel) decreased the

Jan 1, 1948