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By K. P. Anderko

INVESTIGATIONS of the constitution of the system Mg-Ba were first made by Grube and Dietrich1 between 0 and 47 atomic pct Ba and then by Klemm and Dincke1acker" ver the whole composition range. Klemm and Dinckelacker corroborated intermediate phases of the approximate formulas Mg9Ba and Mg4Ba, and showed a third compound to have the formula Mg2Ba, rather than Mg5Ba, as suggested by Grube and Dietrich. Mg2Ba had been previously recognized by Hellner and Laves3 s a MgZn2 (C14) type structure. The main discrepancies between the results of Grube and Dietrich and those of Klemm and Dinckelacker exist, however, in the liquidus temperatures and solidification equilibria of alloys containing more than about 10 atomic pct Ba. The liquidus temperatures measured by Grube and Dietrich, and the outline of the liquidus of Klemm and Dinckelacker are plotted in Fig. 1 on an enlarged scale. Although there are strong reasons to prefer the results of Klemm and Dinckelacker, it was considered worthwhile to remeasure liquidus temperatures of some alloys in the controversial composition range. The alloys, made up from sublimed magnesium and 99.5 wt pct pure barium, were melted in an Armco-iron crucible in a dynamic argon atmosphere. Thermal analyses were taken using a Speedomax auto- matic recorder. Careful stirring during the solidification process was required in order to avoid segregation. After thermal analysis some of the cast alloys were chemically analyzed. The intended and analyzed compositions of the ten alloys prepared are listed in Table I. Since standard barium determinations cannot differentiate between barium present as element and as oxide, both the barium and magnesium contents of the alloys were analyzed. The fact that the sum of barium and magnesium was close to 100 pct in each case was taken as evidence that no excessive oxidation had occurred. The liquidus temperatures measured are listed in Table I and plotted in Fig. 1. As can be seen, they are in good agreement with the liquidus found by Klemm and Dinckelacker. This leaves no doubt that the thermal results of Grube and Dietrich above 10 atomic pct Ba are incorrect. This work was performed under the sponsorship of the Pitman-Dunn Laboratories of Frankford Arsenal, Philadelphia, Contract No. DA-11-022-ORD-1975. Sponsorship and permission to publish by this organization are gratefully acknowledged.

Jan 1, 1958

By M. R. Geer

THE importance of the size composition of coal is reflected in the differ-ence in price of the various sizes of the same coal and in the large number of primary sizes and mixtures of sizes produced by the modern preparation plant to meet the demands of a market that in late years has grown increasingly conscious of coal size. The importance of size in the indus-trial and domestic uses of coal has its counterpart in the role played by size composition in the laboratory examination of coal. Investigation of many of the physical properties of coal in their relation to its uses requires an evaluation of size composition. In fact, the Standards on Coal and Coke issued by the American Society for Testing Materials in 1936 contains 12 standard methods of testing that involve some use of size composition. The screen analysis is, of course, the standard method of determining size composition, whether the material is coarse run-of-mine coal or fine pulverized fuel. Customarily, the analysis is reported as the weight percentages retained on or passing each of a series of testing screens, the number of which may range from only two or three to as many as 20. When more than two or three sieves are used, which is usual, the number of weight percentages obtained is so great as to make the analysis difficult to comprehend. Direct comparison of analyses made on the same sieves is hard, and comparison of analyses covering the same range of sizes but made oil different sieves is next to impossible. Consequently, a frequent purpose of sizing tests-correlation of size with properties and charac-teristics related to size-is rendered difficult.

Jan 1, 1938

By Jeff Steinhoff

INTRODUCTION The Henderson mine utilizes a highly mechanized, continuous, panel-caving, mining system to extract ore from a deep, massive, molybdenite deposit. The mine is located 80.5 km (50 miles) west of Denver, Colorado. The mine surface facilities are located 3,170 m (10,400 feet) above sea level in a steep valley on the eastern side of the Continental Divide. Milling facilities are 24 km (15 miles) west on the western side of the divide at an elevation of 2,804 m (9,200 feet) above sea level. The ore- body is located approximately 3,000 feet south of the valley under Red Mountain. Access to the ore for men and materials is through a 915 m (3,000-foot) deep, 8.5 m (28-foot) diameter, vertical, concrete-lined, service shaft. Access from the mill is through a 15.5 km (9.6-mile) rail haulage tunnel. The mine is ventilated through an additional intake shaft and two exhaust shafts. Mine production at this time is 27,255 mtpd (30,000 stpd). The mine ventilation system supplies 1,038 cubic meters per second (2.2 million cfm) through approximately 60 miles of drifting or 2.7 tons of air per ton of ore mined. There are 130 fans in the mine in fixed locations and in vent lines with 6,900 connected horsepower in the mine. MINING METHOD AND LAYOUT The orebody is divided vertically into two major zones. The upper zone is the 8100-level production area. The bottom zone is the 7700- level production area which is in the early development stage. The rail haulage level at 7500 feet is common to both production zones. Each mine production zone consists of five associated sublevels. The cave undercut level is 16.8 m (55 feet) above the production level. Two boundary cutoff levels are located 44.2 m (145 feet) and 62.5 m (205 feet) respectively above the production level. The fresh-air level is positioned 15.2 m (50 feet) below the production level, and the exhaust vent level is 19.8 m (65 feet) below the production level. Horizontally, each production zone is divided into three panels each, 224 m (800 feet) wide. These panels are caved from south to north. As the caving in one panel nears completion, caving in the adjacent panel is initiated. Development for the caving panels is continuous so that the sublevels above the production level and the production level itself have a combination of development drifting and production-related activities. UNDERGROUND VENTILATION NETWORK The ventilation system is zoned in the same manner as the orebody itself. One major split of 600 cubic meters per second (1,270,000 cfm) ventilates the 8100-level production zone; one split of 100 cubic meters per second (210,000 cfm) ventilates the development of the 7700- level production zone; and one split of 165 cubic meters per second (350,000 cfm) ventilates the 7500 rail haulage level. The haulage tunnel requires an additional 188 cubic meters per second (400,000 cfm) of air. Development-drift ventilation is accomplished by hanging 1.0 m (3.5-foot) diameter steel ducting in the drifts with 40-horsepower, 0.96 m (38-inch) diameter fans supplying 9.4 cubic meters per second (20.000 cfm). The normal maximum length for these systems is 300 m (1,000 feet). The 8100-level production-area ventilation system is especially suited to a high level of mucking activity confined in a small area. Approximately 93 per cent of the mine's total production is transferred from the drawpoints to ore passes in 10 production drifts. The active area in each drift is 300 m (1,000 feet). Twelve 5-cubic-yard LHD units with Cat turbo- charged 170-horsepower engines are assigned to the area. Under these conditions, more than one LHD is assigned to a particular production drift. Adequate ventilation is maintained by making an air change every 97 m (320 feet) along the production drifts. Fresh air is brought into the production drifts from the fresh-air level through 1.37 m (4.5-foot) diameter raises. Air travels south along the production drift to the ore pass where it is exhausted down the ore pass to the exhaust level. The ore pass is followed by another intake which is followed by an ore-pass exhaust. At the south end of the production area, a series of exhaust fans maintain a southerly air- flow through the production level.

Jan 1, 1981

By R. P. Rothwell

AT this our first meeting I beg to call the attention of the members of our Institute to what is certainly a question of the greatest possible importance to the industries we represent; and more particularly to the welfare of the coal and iron trades: I refer to the enormous waste of coal in mining and in its preparation for market, as now practised in our anthracite coal-fields. I t is a fact which many of you can certify to, that we are wasting fully me-half of all the coal in the veins we are now working, and there are cases, which might be mentioned where this proportion, immense as it is, has been greatly exceeded. The popular belief that our deposits of anthracite coal are "practically inexhaustible," is very far from being founded on facts. On the contrary, supposing the present rate of increase in the production to be maintained, it is probable that a maximum yield for our coal-fields would be attained in about fifteen years ; there are in fact some of our smaller basins where the principal veins are nearly exhausted and the production has already attained its maximum. There are others, again, where the depth attained-and it must be borne in mind that we commenced working at the surface outcrop moderate though it is (scarcely ever reaching eight hundred feet), has so increased the cost of mining as to imperil the capital invested, and though it requires but little foresight to predict the time when the difficulties of mining will be very considerably increased by reason of the greater depth, greater quantity of water, fire-damp, and other fruitful sources of cost, yet, ignoring these recognized facts, we continue to waste more than one-half of the coal where it is obtained with the least danger, expense, and trouble, that is, near the surface-either above water-level or at a moderate depth . below it. The inspection of almost any accurate mine map will satisfy you that, on an average, we leave fully one-third. of the coal in the ground as pillars to support the roof; in some places more than one-third thus remains, while it is very rare that less than one quarter is left. In almost every case this coal is absolutely lost, for it will very rarely pay to reopen gangways and lay roads to it after that portion of the mine has been abandoned. Is it necessary in all cases to support the roof, and where it is, may not some artificial

The selection of a particular machine for production drilling is the most critical drill evaluation the pit engineer is called upon to make. It is a true engineering design problem requiring value judgments. Generally, the procedure follows these steps: (1) Determine and specify the conditions under which the machine will be used, including such service factors as labor, site, weather, etc. (2) State the objectives for the rock-breakage phases of the production cycle of operations--considering excavation and haulage restrictions, crushing capacity, production quota, and pit geometry-in terms of tonnage, fragmentation, throw, etc. (3) On the basis of blasting requirements, design the drill-hole pattern (hole size and depth, in- clination, burden, spacing, etc.) (4) Determine the drillability factors and select the drilling methods which appear feasible for the kind of rock anticipated. (5) Specify the operating variables for each system under study, considering drill, rod, bit and circulation fluid factors. (6) Estimate and compare performance parameters, including costs. Major cost items are bits, drill depreciation, labor, maintenance, power and fluids. (7) Select the drilling system that best satisfies all requirements with the lowest over-all cost.

Jan 10, 1967

By S. I. Liu

THE effects of strain histories, consisting of several prestrains in opposite directions, on the flow and fracturing characteristics of metals determined in a final test have been investigated previously.'" Such strain histories vary from a single prestrain in tension or compression to the number of strain cycles leading to fatigue failure.' Between these two limits, various strain histories could be visualized and some can be attained readily. One such example is a strain history consisting of a first prestrain in tension of various magnitudes and a second prestrain in compression also of various mag-nitude. In this investigation the strain history was extended to the combination of these two limits, i.e., a single prestrain in tension or compression followed by balanced strain cycles of a selected magnitude e, = k0.12. It was found in this case that only compressive initial prestrains could be investigated with sufficient accuracy.* Such strain histories can be considered also as (low cycle) fatigue tests of a metal subjected to various amounts of cold working. It appears from previous publications that cold working within a wide range of reductions usually raises the un-notched fatigue strength of many metals.'-' This effect, however, has been found to be inconsistent and sometimes reversed for other metals?.7 Furthermore, Ludwik reports that the yield strength of cold-worked specimens may be reduced by cyclic strains, while that of materials without cold work is generally increased in the initial cycles." " In addition to the above tests, a study was made of the effect of intermediate re-solution heat treatment on low cycle fatigue behavior. Such anneals should eliminate the strain hardening induced by the cyclic strains. As far as their effect on the fracture characteristics is concerned, a few previous tests seemed to indicate that the damage resulting from cyclic strains is retained only partially after heat treatment." A further investigation was made in which one specimen was subjected to repeated unbalanced tension cycles in which the strain interval was approximately $ 0.013. This repetition of plastic tension damages the material slightly. This result appears to confirm Ludwik's observations." " Commercial 3/4-in. rod of the aluminum alloy 24ST4 was selected for this investigation. The material was subjected to the following treatments before machining: 1—Re-solution treatment at 920" ± 10°F for 45 min, 2—water quench at room temperature, and 3—aging for four days at room tem-perature.? The treated material then was machined to desired specimen contour and tested. A coolant was used during machining to minimize any artificial aging effects of heating during machining. In some cases the 24ST4 specimens were reheat treated after having been strained. In these cases, the heat treatment was the same as the re-solution heat treatment described above. Procedure As shown in Fig. 1, threaded end specimens with a minimum diameter of 0.266 in. and a contour radius of 2 in. were used for prestraining to an axial compressive strain ranging from 6, = —0.10 to E, = —0.48. For large prestrains, two consecutive corn-

Jan 1, 1952

By Leo J. Vogel, E. D. Hummer

INTRODUCTION by E. D. HUMMER An efficient refuse-disposal system is a necessary part of the modem cleaning plant. The large-scale refuse system and disposal area, engineered for the lifetime of the plant, has become more important due to the greater tonnages of refuse produced from increased quantities of mechanically mined and prepared coal. Laws and increased government regulations have affected the design and operation of disposal systems. THE EFFECT OF MINE AND PLANT OPERATION Refuse disposal starts at the mining face. The mining system, the type of roof support, the machinery used, and the degree of discipline at the face contributes to the amount and type of refuse that must be removed through the disposal system. Economies gained through modification of the mining system or mining machinery may be partially offset by increased dilution of the mined coal with impurities from associated rock formations. In some cases the change in the mining system and equipment has reduced the dilution. In all cases the ratio of impurities that must be removed from the raw-coal feed to the plant not only affects the maintenance and operating costs of the plant, but greatly affects the design of the refuse disposal system. The design of a mine and plant should include the location and the estimated capacity of the disposal areas for the life of the property or plant. Complete evaluation of the disposal system and an estimate of its annual cost of operation over the life of the property or plant should be incorporated in the plant and mine evaluation. Plant design should be as simple as possible, consistent with the efficient production of clean coal. Material transfer points, crushing, screening and pumping operations should be kept to a minimum to reduce degradation and the subsequent increased difficulty and cost of refuse removal. The maintenance of good process control by operating management will result in a minimum of misplaced material in both refuse and clean coal. Reduction of misplaced float material in the plant refuse is an important factor in preventing combustion in the refuse storage pile. THE REFUSE DISPOSAL AREA Combustion in refuse piles occurs when air enters the pile in sufficient quantity to support oxidation. At certain rates of air in-filtration, the heat produced by the oxidation process is not dissipated and temperature rises.

Jan 1, 1968

By John Mitchell, C. C. Hanninen, L. D. Clark

INTRODUCTION Certain rock types can be fractured thermally by spallation and thermal cracking. Spalling occurs in the regions subjected to thermal shock temperatures and cracking takes place in the adjacent cooler regions. The analysis of the heat-affected zones requires that heat transfer calculations be made where the thermal conductivity of the rock is a function of temperature and direction. Generally, accurate thermal conductivity data is not available for the high temperatures and rise rates associated with most thermal rock breaking methods. This paper describes a method for determining the thermal conductivity of an orthotropic material whose conductivity is a linear function of temperature. The necessary data is obtained from one test which makes it especially useful for in situ applications. The properties data are determined through an iterative process of comparing theoretically computed values of temperature for assumed conductivity parameters with data measured in the laboratory. The conductivity of the rock is determined when the results of the assumed parameters yield computed temperatures which agree with those obtained experimentally. Transient temperatures are computed using the explicit numerical method at given points in a three-dimensional semi-infinite solid whose surface is symmetrically heated on a circular spot. The thermal conductivity is assumed to be a linear function K = [a] + Bt; however, if necessary, any reasonable linear or non-linear function might just as well be used. The stability, convergence, node size,

Jan 1, 1971

By Sherwin F. Kelly

The need for geophysicists to adopt a uniform mode of expressing the electrical resistivity of geological formations has been stressed by Dr. A. S. Eve.1 The present paper is to emphasize the point he raised, to bring the weight of disinterested authority to bear on the matter, and to urge the adoption henceforth, save in exceptional cases, of a standard expression for resistivity; all in the hope that the needed reform will thus actually be brought about. The general definition of the resistivity of a material is the resistance in ohms between two opposite faces of a unit cube of that material. No account need be taken here of the cases in which the unit volume is not a cube, and the units of length and cross-section are different, encountered principally when dealing with metallic conductors such as wires. The idea of the resistance of a unit volume of earth or rock has found expression in geophysical literature in a number of forms and with a variety of units. Fortunately, the metric system has been adhered to in most cases, so ordinarily the meter or the centimeter, usually the latter, has been the measure chosen. The resistivity is then the resistance of a cube of the given rock, or soil, one centimeter on an edge. This is far from always being the same thing, however, as the resistance of a cubic centimeter. A long thin wire could contain a cubic centimeter and have a high resistance, whereas the same volume formed into a flat plate would have an entirely different, and quite low, resistance between faces. Consequently, it is not proper to express resistivity in ohms per cubic centimeter, inch, etc., as has been done occasionally by some writers. In an effort to emphasize the idea of cubic volume, some geophysicists, the writer among them, have used the phrase, ohms per centimeter cubed. This is a graphic way of picturing the measure, but is not in favor among highly qualified authorities. In this connection, an appeal was made to the U. S. Bureau of Standards for advice, and the Director, George K. Burgess, kindly replied in detail as to the practice adopted by his Bureau: 'I.., such expressions as 'ohms per centimeter cubed' or any similar expressions are objectionable, because they certainly carry the implication that the resistivity is found by dividing ohms resistance by the volume in cubic centimeters." Actually, of course, resistivity is found

Jan 1, 1932

By F. D. Rosi, C. A. Dube, B. H. Alexander

WHEN a deformed polpcrystalline metal is heated to a sufficiently high temperature, a recrystallized structure develops which consists of small, essentially stress-free grains. This transformation is referred to as primary recrystallization. Under certain conditions, on prolonged annealing of the specimens, a type of grain coarsening occurs in which a small number of grains commence to grow at various points by devouring the neighboring finegrained primary structure, which, in itself, exhibits a marked stability toward normal grain growth. This grain coarsening process has been referred to by Burgers' as "secondary recrystallization" or "exaggerated grain growth." Although much work has been done on recrystallization and growth processes, present knowledge of the factors which determine the occurrence and growth of secondary grains is inadequate for the advancement of a theory which accounts for the observed characteristics of this recrystallization phenomenon. In this respect, it would be important to make measurements of the velocity of growth of these large secondary grains. The early work of Feitnecht2 on aluminum showed that the size of secondary grains increased with temperature and time, suggesting that such measurements would be possible. More recently, Burgers1 found that the growth velocity of these grains in aluminum was constant over a large time interval, as is generally found for the growth velocity of primary grains in a uniformly deformed test piece.3-5 It is reasonable to expect that the determination of the activation energy of the secondary growth process might lead to a better understanding of the underlying atomic mechanism. This investigation, therefore, is one of a series primarily designed to measure the growth velocity of secondary grains at various temperatures for silver, copper, and aluminum. It is hoped that these studies along with those of orientation relationships, will establish a sound basis for theorizing as to the origin and nature of the secondary transformation. Experimental Procedure The metal used in the present investigation was high-purity silver (999.9 fine), which was supplied in the form of cold-rolled plates 0.250 in. thick. These plates were annealed at 500°C for 1 hr and then straight-rolled to a thickness of 0.04 in. in three reductions of approximately 45 pct, followed in each case by an anneal at 450°C for 1/2 hr. The 0.04 in. sheets were given a final reduction of approximately 50 pct, which resulted in a thickness of 0.02 in., thereby insuring two-dimensional crystal growth. Individual specimens of this material about 1x3 cm were used for all heat treatments, unless otherwise noted. The annealing was carried out under an atmosphere of argon in an Inconel tube furnace whose temperature was controlled to within -+2°C. Metallographic examination of the secondary transformation was accomplished by mechanically polishing the specimens on one side to approximately one half their thickness through 2/0 metallographic emery paper, and then electrolytically polishing in a solution of potassium ferrocyanide and sodium cyanide. With a current density of approximately 20 amp per sq dm, the time required to obtain a satisfactory surface varied from 3 to 5 min. Standard potassium bichromate was used to etch the electrolytically polished surfaces. A heavy etch with this solution produced a sharp contrast between the secondary grains and the fine-grained primary structure, darkening the latter while leaving the large secondary grains bright. All area measurements of the secondary crystals were made with a planimeter on enlarged tracings of the grains.

Jan 1, 1953

By Fred Allison

THE magneto-optic method of analysis had its origin in experiments1 which were designed to detect and measure a time lag in the Faraday effect and later to study this time lag as a function of the wave length of the light.2 Faraday discovered in 1845 that a beam of plane polarized light, when passed through a transparent, isotropic medium traversed by a magnetic field of force parallel to the direction of the light, suffers a rotation. The direction of rotation is reversed with the reversal of the magnetic field and its magnitude depends upon the strength of the field, the length of the path of light in the substance and the nature of the substance. This phenomenon of magnetic rotation is known as the Faraday effect. A question which had been of some interest for years was whether this power of a liquid to rotate polarized light is acquired instantly with the application of the magnetic field, that is, whether it is a case of instantaneity or whether there is a very small interval of time which intervenes between the application of the magnetic field and the consequent rotation of the light by the liquid; in other words, whether there is a time lag in the Faraday effect. Efforts to measure such a lag previously had met with failure. It was recognized at the outset' that since a time lag had escaped detection it must be extremely small, if to a measurable extent it exists at all. It became necessary therefore to adopt a timing device capable of making measures of almost infinitesimally small intervals. The velocity of light served as this timing device, a modification of a method originally employed by Abraham and Lemoine3 being adapted to the requirements of the problem. Since light travels about 30 cm. in a billionth of a second and since settings in our observations may be repeated with deviations not exceeding 3 mm.,

Jan 1, 1932

By Henry Phelps Howland

During the last decade no topic has created more interest or received more thought among blast-furnace men than coke. One reason for this is, undoubtedly, the remarkable increase in the use of byproduct coke. Formerly our coke was made at a distance from the furnaces and by a distinct organization. There was little cooperation or development possible under these conditions. Today the situation is greatly changed. Much of our coke is made by the organization operating the furnaces and the coke oven is operated with the primary object of obtaining the best results in the blast furnace. The effect of changing the method of coke manufacture, even in some minor detail, is often felt almost immediately at the blast furnace. There naturally results one of the most valuable assets of the byproduct coke oven, i.e., the possibility of increased blast-furnace efficiency due to close observance and correct operation of the ovens. The results are evident, but often it is difficult correctly to account for them. In attacking the solution of a problem of as much importance as why one coke is better than another, there are facts which should be kept clearly in mind. The author hopes, in this paper, to bring out some points relative to the use of carbon in the blast furnace which will be of value in solving this question. GeÜner's Ideal Working Years ago GrÜner made a statement regarding the use of carbon in the blast furnace, which has been very generally accepted by metallurgical and blast-furnace men. I quote this theory as given by Prof. Richards on p. 248 of Metallurgical Calculations: "From the standpoint of the generation of the maximum quantity of heat in the furnace, GrÜner was right in formulating his dietum of the ideal working of a blast furnace, vie.: All the carbon burnt in the furnaee should be first oxidized at the tuyères to CO and all reduction of oxides above the tuyères should be caused by CO, which thus beeomes Co2."

Jan 1, 1917

By J. S. Kirkaldy, R. J. Brigham

AS part of a program to investigate the diffusive properties of dilute ternary austenites we have made a determination of the effect of chromium on the activity of carbon in ? iron. Samples of pure iron and Fe-Cr alloys were simultaneously equilibrated at 1000°C with atmospheres of controlled CO/CO2 ratio following the procedures of Darken and Gurry 1 and smith.' The 0.020 in. thick samples were analysed for carbon by microcombustion analyses3 to yield a set of isoacti-vity curves. The latter were converted to activity coefficient (standard state at infinite dilution) plots by combination with Smith's data2 for the activity of carbon in pure iron. Table I shows the compositions of the base alloys used while Fig. 1 shows the experimental results superimposed on a set of curves generated by the relation In?C=6.9NC- 10.7 Ne In the notation of wagner4 the interaction parameter is the second coefficient on the right, Sgr = - 10.7 This can be compared with a rather inaccurate value deduced from the transient equilibrium diffusion data of Kirkaldy et al.' as The authors are indebted to the American Iron and Steel Institute for financial support. This material is condensed from a B.Sc. Thesis submitted to Mc-Master University by R.J.B.

Jan 1, 1963

By M. B. Bever, A. K. Jena

In the investigation reported here, calorimetric measurements have been made to determine the heat of formation of the compound AuSb2 and the partial gram-atomic enthalpies of antimony and gold in dilute solution in liquid tin and their enthalpy interaction coefficients. In order to interpret the behavior of antimony and gold in dilute solution in tin, nonzero values are assigned to the energies of bonds between like solute atoms relative to the pure solute state. Sa.mples of the compound AuSb2 were prepared by melting stoichiometric amounts of antimony and gold (99.999 pct pure) in evacuated and sealed Vycor tubes, which were shaken vigorously and quenched in water. The ingots were annealed in evacuated and sealed tubes at approximately 700°K for 2 weeks. Metallo-graphic and X-ray examinations gave no indication of second phases or microsegregation. The samples were added from 273°K to tin-rich solutions in a solution calorimeter1 maintained at 623°K. The calorimeter was calibrated with tin; the heat content of tin, H623°K- H273°K, was taken as 4.08 kcal per g-atom.2 In each calorimetric run at least four additions were made. The total concentration of solutes in the bath at the end of any run did not exceed 1 at. pct. The measured heat effect of an addition was plotted against the average of the atom fractions x of Sb + Au in solution before and after that addition. From this plot the heat effect at infinite dilution h,,, and the composition dependence of the heat effect dh/dx in the composition range 0 < x < 0.01 were obtained. The results for the additions of antimony, mixtures of antimony and gold, and the compound AuSb2 are listed in Table I. The heat effect h on addition of a sample of the compound AuSb2 or a mechanical mixture of antimony and gold at 273°K to a tin-rich solution at 623°K can be expressed as:

Jan 1, 1969

By F. W. C. Boswell, R. L. Cunningham, H. M. Weld

Chemical analysis shows the partition of the magnesium addition between the metal and the graphite is independent of whether the graphite is in the form of nodules or flakes. The bright central spot observed in nodular graphite polished sections is shown to be due to the structure of the nodules. DURING the past few years numerous papers have been published on the production1-7 and heat treatment" of nodular cast irons. However, very little information is available on the structure of the nodules and the reason for their formation. It would appear that an explanation of the phenomenon of nodulation must await a detailed knowledge of the nature of the nodule. Disagreement exists among various workers on such a fundamental point as to whether or not a nodule actually contains a microscopically visible, nongraphitic nucleus.9-11 Work was initiated with a view to obtaining definite information on this and other aspects of the problem. Chemical Separation and Analysis Microscopic appearance had indicated9-11 that a nongraphitic core might exist in nodules (Fig. 1). It appeared that a chemical comparison of separated nodules and flake graphite would be profitable, for if the core is nongraphitic a relatively large amount of certain elements would be found in the nodules as compared to the flake. This comparison could best be made by obtaining nodules and flake from the same melt of cast iron, part of which had been in-noculated with a suitable addition reagent for the formation of nodules. A 300 lb melt was produced. About half of this was innoculated in the ladle with an Fe-Si-Mg alloy to produce nodules and cast into keel blocks; the remaining half of the melt had silicon added in an effort to compensate for the silicon in the innoculant used in the other half of the melt and this was cast into similar keel blocks. The analyses of the resultant irons are given in Table I. Nodules and flake were chemically extracted from 2 kg samples of metal as described in the appendix. While this method dissolved the iron and chromium carbides, it did not dissolve the titanium, vanadium, and molybdenum carbides. These latter carbides could have been removed by nitric acid, but in other experiments it was found that it attacked pools of metallic iron entrapped in the nodules which had, until the nitric acid treatment, been protected from the attack of the sulphuric, hydrochloric, and hydrofluoric acids. It will be shown in the next section that entrapped iron exists in nodules but not in flake. On microscopic examination many tiny irregular particles were observed to accompany both the nodules and the flake. These were identified by tabling a sample of nodules from which a sufficient fraction was obtained for X-ray analysis. A good pattern of titanium carbide was obtained indicating that this was the predominant compound present. Since it was not possible to determine how completely the carbides could be removed mechanically and since their removal from flake is much more difficult, it was decided not to resort to any separation as this would render further comparisons more difficult. The samples were ashed in a combustion

Jan 1, 1953

By Charles E. Johnston

By substituting photoelectric automatic car loading for mechanically induced hydraulic automatic loading and by protecting the process against operational failures, the engineering staff at International Minerals & Chemical Corp.&apos;s Carlsbad potash mine has developed a reliable loading process that operates without an attendant. The photoelectric control, developed primarily because of frequent failures in the hydraulic control system, has not changed the basic loading operation and physical layout. Mine-run ore ranging in size from minus 200 mesh to 30-in. lumps is dumped onto a conveyor belt which discharges into a so-called "flop gate chute" which in turn, feeds the ore into six-ton capacity mine cars. Each car is filled progressively by moving the entire trip forward. When a car is filled 12 in. above its rim, the direction of flow into the car is reversed 180º and filling of the next car back is begun while the entire trip is moved forward again. Once the next car is properly spotted under the front chute, the flow is again directed to the forward position.

Jan 9, 1962

By Jesse Betterton

IN the Parkes process of lead refining, after desilverization has been completed by means of zinc additions, there will remain in the lead from 0.5 to 0.6 per cent zinc. At this stage in the refining practically all impurities are reduced to refined lead specifications except zinc and antimony. All excess of copper and nearly all the arsenic goes out with the silver. Silver remaining should not exceed 0.1 oz. per ton. It thus remains only to remove zinc and antimony to produce refined lead. The older practice was to complete refining by means of oxidation processes in reverberatory furnaces. Essentially this consisted of heating the lead to 1200° to 1300° F. and allowing it to come in contact with a considerable excess of air passing through the furnace, or accom-plish the same result by blowing the lead with steam by means of pipes introduced for that purpose, at the same time allowing air to enter the furnace. Usually some procedure involving alternate oxidation and reduction with fine coal or coke was employed in order to reduce the weight of dross or skim finally removed from the furnace. This practice had many elements of weakness in that the main flow of lead, after the preceding expensive operations had been performed, and the whole refining process nearly completed, was subjected to an operation at a comparatively high temperature, which involved much lead loss if the furnaces were not on bags, and the production of 3 to 5 per cent of the weight of the lead as dross or skim, which had to be resmelted and the lead either refined or produced as antimonial lead. In addition, the operating expense of the process was a considerable item. The, process of dezincing with chlorine was developed as a substitute for the foregoing described oxidation process, and finally resulted in such advantages as nearly equal operating costs, a practical elimination of the lead loss, and the direct production of a very valuable byproduct, zinc chloride, that may be sold as such or retreated for its zinc and chlorine contents, both of which can be reused in the refining processes.

Jan 1, 1933

By C. H. Li, R. D. Carnahan, R. J. Stokes, T. L. Johnston

When silver chloride deforms by pencil glide at temperatures of 26ºand 72°C, grain size has no effect upon the proportional limit and the material necks down to a knife edge under tension. At -196ºC, deformation takes place on fewer slip systems to produce straight slip traces, the flow stress becomes sensitive to grain size and fracture occurs by cleavage from an intergranular source without any reduction in area. THERE are significant differences between the appearance of the slip bands formed in silver chloride and in the alkali halides at room temperature. The bands in sodium chloride, e.g., are fine and geometrically straight, whereas in silver chloride they are coarse and wavy.&apos; Such differences in the slip behavior prompted us to inquire what the effects of grain boundaries might be on the mechanical properties of these solids. As an introduction to a general study of ionic poly-crystal behavior, we have investigated the effect of grain size on the stress-strain relationships in silver chloride at different temperatures. We have chosen silver chloride chiefly because polycrystals can be easily prepared free from cracks and macroscopic voids by conventional techniques. Moreover, as we shall see, the slip behavior can be readily modified by changing the temperature. It will be shown that the effect of grain size on the yield stress of silver chloride is strikingly dependent upon the temperature of deformation and that this temperature effect comes about through a change in the number of slip systems which operate at the onset of plastic flow. EXPERIMENTAL PROCEDURES Polycrystalline tensile specimens were made by recrystallizing cold-rolled AgCl sheet. Thirty-gram quantities of AR grade silver chloride powder were melted in a low heat capacity furnace, cast into cylindrical forms 3/4 in. in dim and approximately 1 in. in length and then extruded with a 16:l reduction ratio into rods 3/16 in. in dim. Extrusion temperatures were varied from room temperature to 350°C. In order to impart various degrees of cold&apos;work to the extruded rod. All extrusions were then cross rolled to sheet 0.040 in. in thickness at room temperature and machined to form tensile specimens having a gage length of 0.750 in. with a width of either 0.250 or 0.145 in.* Annealing was carried out in an air oven at 200°C for periods up to 20 hr. specimens having average grain diameters ranging from 0.024 to 0.40 mm were prepared by varying the extrusion temperature and annealing time in the conventional manner. To obtain specimens with grain sizes larger than 0.40mm, lightly deformed thin prisims of silver chloride single crystals (supplied by Harshaw Chemical Co.) were recrystallized. Grain size determinations were made by an intercept method and revealed that the grains were equi-axed. Laue back-reflection X-ray patterns showed little tendency to form a preferred orientation texture. Specimens which were to be used for the slip band studies were polished by immersion in fresh boiling concentrated NH,OH for 60 sec. They were then etch polished for 10 sec on each side by hand lapping on a stationary velveteen lap wetted with a dilute Hypo solution (5 g of Kodak Acid Fixer in 100 cc H20). Finally the specimens were rinsed in warm water and alcohol, blown dry and carefully mounted in the tensile machine for testing. Tensile tests were performed at three temperatures, namely, 26º, -72º, and -196°C. A screw-driven hard tensile machine was used at a constant strain rate of 0.006 in. per in. per min. Loads were determined with a light- proof ring fitted with A-18 strain gages and extensions were measured with a 0.0001 in. dial gage. Constant level baths of alcohol with dry ice and liquid nitrogen were used to maintain the two lower test temperatures. EXPERIMENTAL RESULTS Reproducibility—In view of the light sensitivity and reactive nature of silver chloride, it was necessary to determine whether reproducibility depended upon either exposure to light or to the condition of the surface. However, during the course of the work it became clear that variation of laboratory lighting

Jan 1, 1962

By R. E. Wade

THE Butte, Anaconda & Pacific electrification is of peculiar interest, in an incidental way, to the entire mining fraternity, and especially the engineering branch, not only in this great Northwest country but throughout the entire Western mining country, which is rich in minerals and water power. It is probably fair to say that the prime object of mining and smelting engineering is to reduce the cost of production, and in the case in hand the relative location of mines and smelter brings the cost of transportation of ore to the front. While some might assign this item to the railway people, it is inseparably a part and parcel of the general scheme of the production of metal from the ore. In considering the electrification of the B., A. &. P. railway the mining and smelting engineer, above all, wants to know what saving will be effected by the change in motive power. Unfortunately the new power has been in use during a very limited period and on the Smelter Hill section of the road only, and no reliable data, from actual service conditions, can be obtained until the entire electrified section has been in service for a reasonable length of time. Any remarks on this subject at the present time, therefore, must of necessity be general in nature and more or less descriptive. The railway, electrical and mining worlds and the country in general are indebted to the B., A. & P. railway for the pioneer work it is now doing, as it undoubtedly marks the beginning of the electrification, within the comparatively near future, of all transcontinental roads between the Missouri river and the Pacific coast. Together with this goes the development of an untold wealth in water power and all natural resources within the same territory. For the first time in the history of electric railway work, in this country, direct current at a potential of 2,400 volts is being used. With this potential there is a great saving in transmission of energy to the locomotives and there are only two substations, one at. either end and 26 miles apart, for supplying the power for the movement of trains of heavy tonnage over the main line and for switching movements at terminal and intermediate yards, and comparatively little feeder copper is required.

Jan 11, 1913

By J. R. Campbell

IN order to discuss the subject intelligently, it will be necessary to touch briefly on the forms in which sulphur is supposed to exist in coking coal to be carbonized in beehive or byproduct ovens. Sulphur is known to exist in coal as sulphides and sulphates, as can be determined experimentally. Then there is the so-called organic sulphur, i.e., sulphur in combination with the carbon, hydrogen, and oxygen of the coal, about which much has been written, but nothing definitely proven in an experimental way. In fact, so far as we know, there never has been developed a satisfactory and conclusive method in the laboratory for the direct determination of organic sulphur. For most coking coals, it can be safely assumed that the preponderance of sulphur is in the form of pyrite (FeS2). When exposed to comparatively low temperatures during the coking process, it loses its sulphur according to the following chemical reaction: 7FeS2 = Fe7S8 + 6S. It will be seen, therefore, that six out of the 14 atoms of sulphur (or 42.8 per cent.) are expelled, or volatilized. Fe7S8 remains in the coke as pyrrhotite, or magnetic sulphide of iron; this accounts for the highly magnetic properties of the powdered coke. A more commonly accepted reaction is FeS2 = FeS + S, in which the straight sulphide of iron is produced with 50 per cent. volatilization of sulphur. In the beehive methods of coke making, air is introduced in sufficient amounts to carry on the distilling and the coking processes, and the sulphur is oxidized along with the other volatile products: First, into sulphur dioxide (SO2), known by the pungent and suffocating odoremitted from the trunnel head; second, into sulphuric anhydride (SO3); and, finally, into sulphuric acid, as it comes into contact with the air and moisture. In a properly regulated draft on a beehive oven, there is never complete combustion of the gases. In other words, the coking process should be carried on in a reducing atmosphere, and a low-grade producer gas kept issuing from the trunnel head. For coals of about 30 per cent. volatile matter, the ratio of air to gas is 3 ½ to 1. In complete combustion, the ratio is 6 to 1, producing an extremely high temperature

Jan 1, 1916