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By C. G. Fink, G. L. Putnam

The dissolution of gold by cyanide solutions was studied by determining the time required for the solvents to dissolve gold leaf. Minute traces, even 0.5 ppm, of sulphide ion retard the dissolution of gold, and this behavior cannot be accounted for by the presently accepted hypotheses involving oxygen-depletion or thiocyanate formation. On the other hand, traces of the salts of lead, bismuth, mercury, and thallium considerably accelerate the dissolution. The beneficial effect of lead is dependent upon the pH and cyanide ion concentration of the solution. ALTHOUGH cyanide solvents are used very ex-x*. tensively for the recovery of gold from ores and concentrates, the normal rate of solution of the precious metal as determined by Barsky, Swainson, and Hedley,1 and others2 is less than about 3 mg per sq cm of exposed gold surface per hour, corresponding to a corrosion depth of approximately 1.5 microns (0.00006 in.) per hr. Although the slow rate may be increased to some extent in certain cases, as, for example, when the the gold is partially imbedded in iron pyrites,3 the leaching operation is generally the most time-consuming step of the cyanide process. Frequently it is necessary for the leaching solvents to remain in contact with the ores for 50 to 200 hr in order to dissolve all of the gold particles.' Not only does the slow solution rate cause some increase in the capital costs, but the cyanide consumption also may be increased because of vaporization of hydrocyanic acid and secondary reactions with minerals such as malachite (CuCO3 • Cu(OH)2), and pyrrhotite, which often are associated with gold. It is not surprising that numerous attempts have been made to accelerate the dissolution of gold in cyanide solutions. Experimental Gold Leaf Test: The gold leaf test was apparently originated by M. Faraday, who, by means of it, discovered that dilute cyanide solutions would dissolve gold readily in the presence of oxygen. The usefulness and reliability of the test have been discussed elsewhere.5,6 Essentially, the method consists of determining the time required for 5-ml portions of cyanide solvents to dissolve squares of 23 carat gold leaf, 0.5 cm on edge, when shaken in 10-ml test tubes with the solvents. The weight of gold leaf per test was 0.051 mg as determined by the direct weighing of a leaf 8.6x8.6 cm (weight = 15 mg) and measuring the area of the test sample (0.25 sq cm). Gold leaf is of remarkably uniform thickness.6 Since gold leaf, like native gold, contains small amounts of copper and silver, the results are comparable with those one might expect in cyaniding gold ores. As six or more tests can be made simul- taneously, the relative efficiencies of the solvents can be evaluated readily. All dilute solutions of lead, bismuth, thallium and mercury salts were prepared by the ordinary dilution method used by chemists. Our C.P. sodium cyanide and buffering reagents had no detectable amounts of either sulphide ion or heavy metals. Soluble Sulphides in the Cyanide Process: The action of soluble sulphides is of interest in the study of the dissolution of gold in cyanide solutions. That gold ores which contain compounds of arsenic and antimony often are not amenable to direct cyanide treatment is well known, and it said that arsenic and antimony are "cyanicides."7 It is commonly believed that part of the sulphur content of such ores is soluble in alkaline solutions to give sulphide ion, which reacts with the free cyanide content of the leaching solutions to form inert thio-cyanates or reacts with the dissolved oxygen to form sulphites or sulphates.8 A purpose of this paper is to demonstrate that sulphide ion may behave in a manner that cannot be explained by either the oxygen-depletion or thiocyanate formation theories and to prove directly that concentrations of sulphide ion which cannot be detected even by colorimetric methods may retard the dissolution of gold.

Jan 1, 1951

By C. G. Fink, G. L. Putnam

The dissolution of gold by cyanide solutions was studied by determining the time required for the solvents to dissolve gold leaf. Minute traces, even 0.5 ppm, of sulphide ion retard the dissolution of gold, and this behavior cannot be accounted for by the presently accepted hypotheses involving oxygen-depletion or thiocyanate formation. On the other hand, traces of the salts of lead, bismuth, mercury, and thallium considerably accelerate the dissolution. The beneficial effect of lead is dependent upon the pH and cyanide ion concentration of the solution. ALTHOUGH cyanide solvents are used very ex-x*. tensively for the recovery of gold from ores and concentrates, the normal rate of solution of the precious metal as determined by Barsky, Swainson, and Hedley,1 and others2 is less than about 3 mg per sq cm of exposed gold surface per hour, corresponding to a corrosion depth of approximately 1.5 microns (0.00006 in.) per hr. Although the slow rate may be increased to some extent in certain cases, as, for example, when the the gold is partially imbedded in iron pyrites,3 the leaching operation is generally the most time-consuming step of the cyanide process. Frequently it is necessary for the leaching solvents to remain in contact with the ores for 50 to 200 hr in order to dissolve all of the gold particles.' Not only does the slow solution rate cause some increase in the capital costs, but the cyanide consumption also may be increased because of vaporization of hydrocyanic acid and secondary reactions with minerals such as malachite (CuCO3 • Cu(OH)2), and pyrrhotite, which often are associated with gold. It is not surprising that numerous attempts have been made to accelerate the dissolution of gold in cyanide solutions. Experimental Gold Leaf Test: The gold leaf test was apparently originated by M. Faraday, who, by means of it, discovered that dilute cyanide solutions would dissolve gold readily in the presence of oxygen. The usefulness and reliability of the test have been discussed elsewhere.5,6 Essentially, the method consists of determining the time required for 5-ml portions of cyanide solvents to dissolve squares of 23 carat gold leaf, 0.5 cm on edge, when shaken in 10-ml test tubes with the solvents. The weight of gold leaf per test was 0.051 mg as determined by the direct weighing of a leaf 8.6x8.6 cm (weight = 15 mg) and measuring the area of the test sample (0.25 sq cm). Gold leaf is of remarkably uniform thickness.6 Since gold leaf, like native gold, contains small amounts of copper and silver, the results are comparable with those one might expect in cyaniding gold ores. As six or more tests can be made simul- taneously, the relative efficiencies of the solvents can be evaluated readily. All dilute solutions of lead, bismuth, thallium and mercury salts were prepared by the ordinary dilution method used by chemists. Our C.P. sodium cyanide and buffering reagents had no detectable amounts of either sulphide ion or heavy metals. Soluble Sulphides in the Cyanide Process: The action of soluble sulphides is of interest in the study of the dissolution of gold in cyanide solutions. That gold ores which contain compounds of arsenic and antimony often are not amenable to direct cyanide treatment is well known, and it said that arsenic and antimony are "cyanicides."7 It is commonly believed that part of the sulphur content of such ores is soluble in alkaline solutions to give sulphide ion, which reacts with the free cyanide content of the leaching solutions to form inert thio-cyanates or reacts with the dissolved oxygen to form sulphites or sulphates.8 A purpose of this paper is to demonstrate that sulphide ion may behave in a manner that cannot be explained by either the oxygen-depletion or thiocyanate formation theories and to prove directly that concentrations of sulphide ion which cannot be detected even by colorimetric methods may retard the dissolution of gold.

Jan 1, 1951

By P. G. Shewmon, F. N. Rhines

THE determination of the self-diffusion coefficient of magnesium has been made possible recently by discovery1-1 of a radioactive isotope, Mg28 having a half-life of 21.3 hr,1 and subject to manufacture in useful quantity. In the present research this material was condensed from the vapor phase upon a surface of high purity magnesium. The progress of diffusion of the tracer atoms into polycrystalline magnesium was followed by machining layers and measuring the change in the intensity of radiation as a function of the distance of each layer from the surface. The self-diffusion coefficient was found to be 2.1 X 10-8 sq cm per sec at 627°C, 3.6 X 10-9 sq cm per sec at 551°C, and 4.4 X 10-10sq cm per sec at 468°C; the activation energy is about 32,000 cal per mol. Experimental Procedure Since there was no other published measurement of a diffusion velocity in any magnesium-base material, is was necessary to employ a number of new experimental techniques. The short half-life of Mg28 made it necessary to complete the entire experimental procedure within three or four days. This meant that the work had to be done where a cyclotron was readily accessible and that all operations, prior to the diffusion heat treatment, had to be so designed as to minimize their time requirements. Unusual problems were imposed also by the chemical reactivity of magnesium, its high vapor pressure, and the fact that no satisfactory method for elec-trodepositing magnesium on magnesium is presently available. Finally, the machining and handling of the easily air-borne radioactive-magnesium chips involved certain health hazards, resulting in the need for further experimental restrictions. Preparation of Mg28 The Mg28 was produced in the Carnegie Institute of Technology syncrocyclotron by the neutron spallation of chlorine.5 his involved bombarding a 2 gram crystal of high purity NaCl with a beam of 350 mev protons for a period of 2 hr, after which the crystal was dissolved in warm water and the Mg28 was concentrated and purified by chemical means (see Appendix). About 50 microcuries of Mg28 thus were obtained in the form of magnesium oxinate (8 hydroxyquin-olatc?), which was ignited in air to produce MgO. This in turn was reduced to magnesium metal vapor, by the method of Russell, Taylor, and Cooper," in the vacuum apparatus shown schematically in Fig. 1. Here the essential part is a tantalum ribbon, slightly dished to receive the MgO. The ribbon, pre- viously outgassed at high temperature, is heated to about 1700°C by passing an electric current through it, whereupon tantalum oxide is formed, magnesium vapor is released almost instantaneously, and condensed partly upon the diffusion sample. Diffusion-Sample Preparation: Hot-extruded magnesium rod, 21/32 in. round was used in making the diffusion specimens. The magnesium analyzed as follows: 0.004 pct Al, 0.027 pct Fe, 0.040 pct Mn, 0.0004 pct Cu, 0.0002 pct Ni, and less than 0.01 pct Ca, 0.0004 pct Pb, 0.0011 pct Si, 0.001 pct Sn, and 0.001 pct Zn. A brief study of the crystal texture of this material revealed a sharp fiber texture with the (001) plane roughly parallel to the extrusion axis. Cylindrical samples 1/2 in. long by 5/8 in. were machined from this rod, the end faces dressed on 3/0 emery, and lightly etched with 20 pct HC1 in water. These samples then were annealed for at least twice the intended time of diffusion, at the intended diffusion temperature, in order to stabilize the grain structure at about 1 mm average diameter. The annealing treatments were conducted in argon in the same apparatus and in the same manner as the subsequent diffusion treatments, which will be described presently. Thus, a strain-free plane surface was produced, but there remained a layer of MgO which had largely to be removed before the layer of Mg28 was deposited. Most of this layer was taken off by two light passes over 3/0 emery paper. The balance of the oxide and a thin layer of metal were then removed by etching 5 to 10 min in 4 pct nital (4 pct HNO3 and 96 pct ethyl alcohol) made with absolute alcohol. There followed immediately three quick rinses in: 1-49 1/2 pct methanol, 49 1/2 pct acetone, and 1 pct formic acid, 2-50 pct methanol and 50 pct acetone, and 3-pure benzene. This procedure is essentially that of Sturkey.7 The resulting surface, which was of almost elec-tropolished brightness, remained plane and was free of cold work. It could be kept clean by storing under benzene, or in a desiccator; short exposure

Jan 1, 1955

By V. B. Kurfman

The temperature, strain rate, and orientation deDendence of defbrnzation of single-crystal MgAg has been examined. The crystals exhibit a tendency to single glide and little or no hardening at 25°C for many orientations. A much higher hardening rate is observed when multiple glide occurs, such as can be initiated by surface defects. The tendency for easy glide becomes less dependent on surface preparation and orientation as T — 100°C and bars so tested often fail after one-dimensional necking-. At T > 200°C (transition temperature for single-crystal notch sensitivity and poly crystalline ductility) single glide diminishes and two-dirnensionul necking begins. The crystals do not strictly obey a critical resolved shear stress law, but show the influence of {loo) cracks in determining the slip mode. The results are correlated with the difficulty of sciperdzslocation intersection and semibrittle behavior of this compound in single-crystal and poly crystalline form. Comparisons are made with the slip selection mode observed in tungsten, with the reported observations of easy glide in bee metals. and with the mechanical behavior of poly crystalline MgAg. PREVIOUS work on tensile deformation of polycrys-talline MgAgl and bending deformation of single-crystal MgAg2 has shown that the compound is semi-brittle (i.e., notch and grain boundary brittle). If this semibrittleness is supposed to result from the difficulty of multiple glide (associated with the problems of superdislocation intersection) one might expect single crystals deformed in tension to show pronounced single glide and strong orientation dependence of hardening rate. These experiments were done to examine this supposition and to study the tensile deformation of a highly ordered system which may be considered bcc if the difference between the two kinds of atoms is ignored (actual structure: CsC1). EXPERIMENTAL Single-crystal ingots were grown by directional freezing as previously described.' These ingots were sliced into a by a by 2 in, rectangular bars by electric discharge machining, then round tensile bars were conventionally machined to 1/8-in.-diam by 1-in.-long reduced section. The bars were typically tested without an anneal because of the problem of magnesium vapor loss and they were typically tested as mechanically polished. The analyses are within the same limits as those reported earlier; i.e., the average composition for each specimen is within 0.5 at. pct of stoichiometry, while the total range from end to end in a given specimen varies from 0.7 to 1.4 at, pct. There has been no indication in the results of any variation in slip or fracture mode attributable to the composition fluctuations. The slip systems were determined by two-surface analysis of the bars after testing to failure at room temperature. Single glide was so dominant that there was little difficulty in identification of the dominant slip system even though the tensile elongation to failure often approached 7 to 8 pct in room-tempera- ture tests. Elevated-temperature testing was done in a silicone oil bath and low-temperature testing was done in liquid Np or a dry-ice bath. All stress measurements are reported as engineering stress unless otherwise specified, and crosshead travel is used as the strain measurement. RESULTS The tendency toward single glide is best seen in the pictures, Figs. 1, 2, and 3, which depict deformation at fracture as a function of test temperature. While it is possible to find regions of secondary slip by careful microscopy, such regions are very small. The development of a ribbon-shaped configuration from an initially round section bar pulled at 100°C is typical, occurred by single glide, and illustrates the degree to which such glide continues. At temperatures =100°C the bars typically show elongation of 20 to 50 pct by predominently single glide. Despite the large elongation, fracture even at 150°C occurs in a brittle mode, Fig. 2, in the sense that it is an abrupt failure which shows no discernible necking in the second dimension of the bar's cross section (i.e., there is no appreciable action of any slip modes which would decrease the broad dimension of the cross section). Near 200°C the fracture mode changes slightly. Although most of the sample extension is by single glide, after the bar develops the characteristic ribbon shape it begins to neck in the second (i.e., broad) cross-sectional dimension. The bar becomes very thin in the "necked down" region, Fig. 3, and the reduction in area approaches 100 pct. Often there oc-

Jan 1, 1967

By H. I. Aaronson, C. Laird

Although the past results of X-ray experiments indicate that the broad faces of 0' plates are coherent with their matrix, dislocations lying in arrays have frequently been observed at these boundaries by transmission electron microscopy. Critical experiments employing the latter technique have been carried out in order to determine the origin of these dislocations. It is concluded that 8' plates are essentially coherent with the matrix at their broad faces throughout the aging temperature/time envelope studied. Virtually all of the dislocation arrays observed are deduced to have been formed by plastic deformation accompanying transformation. The proportion of dislocations arising from convexity of the plates is shown to be negligible by comparison with that from plastic deformation. At the higher aging temperatures, a[001] dislocations appeared in moderate numbers. These dislocations were traced directly, however, to the ledgewise dissolution of 0' occasioned by the formation nearby of 0 crystals. On the other hand, since there is a parametric difference normal to the broad faces of the ?' plates, mismatch dislocations do form at their edges. A previous conclusion that these dislocations have Burgers vectors of type a[001] was confirmed directly. The edges of 0' plates were observed to develop octagonal shapes when growing, but circular shapes during dissolution. 1 HIS paper presents the results of an investigation of the interfacial structures of plates of the transitional phase, 8', formed in an A1-4 pct Cu alloy. In a companion paper, Part 11, the effects of these structures upon the migration kinetics of a:?f boundaries are reported. This work is pa.rt of a general program designed to establish the basis of precipitate morphology. The present authors in Al-Ag,1 and whitton2 previously in U-C alloys, have used transmission electron microscopy to examine directly the vander Merwe3-6 networks of dislocations anticipated7 to compensate the small amount of lattice misfit normally founda at the broad faces of Widmanstatten plates. Since the broad faces of 0' plates are considered to be perfectly coherent with the corresponding habit planes in the a matrix,' no dislocations should be present at these faces. Many reports have been published, however, giving evidence to the contrary.10-18 The primary objective of this investigation was therefore to ascertain the nature of these dislocation structures. An attempt to do this is described in the first three sections of this paper. Inspection of the matching of the a and 8 ' lattices at the orientations of the 0:0' boundary corresponding to the edges of 0' plates raises the possibility that these edges may be made up of rather closely spaced edge- type misfit dislocations oriented so as to be sessile with respect to the lengthening or shortening of the plates. Since this structure should severely inhibit migration of the plate edges (Ref. 7, Part II), a situation not originally anticipated,' an experimental determination of the interfacial structure of the edges of 8' plates was clearly in order, and is reported in Section III. Those aspects of the experimental procedure applicable to both Parts I and I1 are presented in the next section. Specific procedures applicable to individual aspects of each investigation, and also the relevant surveys of the literature, are then individually reported in the appropriate sections. I) GENERAL EXPERIMENTAL PROCEDURE The material used in both parts of these studies was the same as that of a previous investigation:" strips of A1-3.93 pct Cu, 0.009 in. thick, prepared as before, solution-annealed at 548°C for 6 hr, and quenched. Details of subsequent aging, and in some cases deformation treatments, are given in the Experimental Procedure sections of the individual parts of both papers. Specimens of the heat-treated strips were electro-thinned as beforeLg and examined in a Philips EM 200 microscope equipped with a goniometer stage. A commercial hot stage, of the grid-heater type and capable of * 30-deg tilt about one axis in the plane of the specimen, was also used for kinetic studies. The usual precaution of calibrating for the additional heat supplied by the electron beam was taken.19 A 16-mm cine cam-I era mounted outside the viewing window was frequently used to record the transformations. Conventional selected-area diffraction and dark-field viewing techniques were used to identify the precipitates in the foils. Normal bright-field images corresponding to two-beam diffracting conditions or dark-field images were employed to characterize the dislocations observed at the interfaces of the precipitates. The application of these techniques to the study of an interphase boundary, and the interpretation of the images,20'21 has been fully described in a previous paper.'

Jan 1, 1969

By Robert A. Colten, Frederick C. Hull, Robert F. Mehl

It is known that pearlite forms from austenite by a process of nucleation and growth, and that the rate of formation of pearlite may be described by a rate of nucleation and a rate of growth.l,2 The manner in which these two quantities operate to produce the observed isothermal reaction curve has been analyzed in a paper published in I939.3 Two types of process were described: (I) that in which nucleation occurs at the grain boundaries of the parent phase, with the nodules growing toward the centers of the grains, initially as hemispheres and then after impingement as roughly radial-columnar structures; and (2) that in which nucleation occurs generally, with the nodules growing to full spheres until impingement with other growing nodules restricts growth. Both cases are met in the formation of pearlite from austenite (Fig. I);2,4 the former, occurring near the knee of the S-curve, has been named "grain-boundary transformation," and the latter, occurring at temperatures near Ael, has been named "group-nodule transformation." In developing our knowledge of the factors that determine the rate of formation of pearlite from austenite, it is important1,2 to obtain data on the rate of nucleation and the rate of growth of pearlite. Such data, when obtained near the knee of the S-curve, will furnish information of use in understanding the factors that affect hardenabil-ity; when obtained at higher temperatures, such data will be useful in understanding the behavior of steels that during heat-treatment react to pearlite.2 It is the purpose of this paper to present data on the rate of nucleation and the rate of growth of pearlite in dependence upon composition, grain size, temperature, and the degree of homogeneity of the parent austenite. The studies are restricted chiefly to plain carbon commercial steels." Previous Data on Rate of nucleation and rate of Growth Data on the rate of nucleation and the rate of growth of pearlite have been published by Scheil and Lange-Weise,5 by Mirkin and Blanter,6,7 by Meh1,l and by Dorn, De Garmo and Flanigan.2, It is difficult to interpret the data, for in most cases too little information is given on austenitizing temperatures and times, the state of homogeneity of the austenite (presence or absence of carbide, etc.), austenite grain size, and mode of formation of pearlite at the temperatures given. Except for the approximations of Mehl and the work of Dorn, austenite grain size was disregarded, though it is well known

Jan 1, 1942

By Robert A. Colten, Frederick C. Hull, Robert F. Mehl

It is known that pearlite forms from austenite by a process of nucleation and growth, and that the rate of formation of pearlite may be described by a rate of nucleation and a rate of growth.l,2 The manner in which these two quantities operate to produce the observed isothermal reaction curve has been analyzed in a paper published in I939.3 Two types of process were described: (I) that in which nucleation occurs at the grain boundaries of the parent phase, with the nodules growing toward the centers of the grains, initially as hemispheres and then after impingement as roughly radial-columnar structures; and (2) that in which nucleation occurs generally, with the nodules growing to full spheres until impingement with other growing nodules restricts growth. Both cases are met in the formation of pearlite from austenite (Fig. I);2,4 the former, occurring near the knee of the S-curve, has been named "grain-boundary transformation," and the latter, occurring at temperatures near Ael, has been named "group-nodule transformation." In developing our knowledge of the factors that determine the rate of formation of pearlite from austenite, it is important1,2 to obtain data on the rate of nucleation and the rate of growth of pearlite. Such data, when obtained near the knee of the S-curve, will furnish information of use in understanding the factors that affect hardenabil-ity; when obtained at higher temperatures, such data will be useful in understanding the behavior of steels that during heat-treatment react to pearlite.2 It is the purpose of this paper to present data on the rate of nucleation and the rate of growth of pearlite in dependence upon composition, grain size, temperature, and the degree of homogeneity of the parent austenite. The studies are restricted chiefly to plain carbon commercial steels." Previous Data on Rate of nucleation and rate of Growth Data on the rate of nucleation and the rate of growth of pearlite have been published by Scheil and Lange-Weise,5 by Mirkin and Blanter,6,7 by Meh1,l and by Dorn, De Garmo and Flanigan.2, It is difficult to interpret the data, for in most cases too little information is given on austenitizing temperatures and times, the state of homogeneity of the austenite (presence or absence of carbide, etc.), austenite grain size, and mode of formation of pearlite at the temperatures given. Except for the approximations of Mehl and the work of Dorn, austenite grain size was disregarded, though it is well known

Jan 1, 1942

By Walter Knox, J. D. Doherty

In 1939, at the request of The Koppers Coal Co., the Koppers Company Research Department established a Stoker Coal Research Laboratory for the purpose of investigating the performance characteristics of stoker coals in typical domestic stoker installations. Fig. I is a photograph of part of the laboratory. The material in this paper deals with the performance of one coal in a single installation. Presented are details of this installation, mechanical characteristics of the stoker, and the results of a series of burning tests in which adjustments were systematically varied. Summary OF Results Mechanical Features I. A large percentage of the total power consumed by a domestic stoker is used to operate the gear case and motor (Fig. 3)- 2' The power consumption per ton goes down with an increase in rate of coal feed (Fig. 4). 3. The automatic air control in this stoker showed advantages Over manual control (Fig. 5). 4. The air velocities through the retort tuyeres were uniform around the retort, but varied between the rows of tuyere rings (Fig. 6). 5. The crushing of coal with this stoker increased the minus 1/8-in. from 11.6 per cent to 19.2 per cent. There was a notice- able segregation of coal as it came from the retort and distribution was not uniform (Fig. 7). 6. Measurement of air entering over the fire showed that with some not uncommon adjustments this air may be a high percentage of the total air supply (Fig. 8 and Table I). Operating Performance ,. The, removal of clinker was shown to give a period of poor performance characterized by a high fuel-bed resistance and a high percentage of carbon monoxide in the flue gases (Fig. 10). 2. The fuel-bed resistance was found to decrease as the fuel bed built up in a normal response to heat demand after a prolonged hold-fire (Fig. II). 3. with intermittent operation the burning rate was almost always less than the feeding rate, The wide variations were found to be greatly dependent upon the length of the On period and the length of the intervals between On periods (Fig. 12). 4. The maintenance of a low furnace draft led to less fly ash, better clinker formation, and higher efficiencies (Fig. 13). 5. The maintenance of a minimum of excess air also resulted in less fly ash, better clinker formation, and higher efficiencies (Fig. 14). Guiding Principles In planning the investigations, the following principles were adopted: Selection of Heating Equipment.—Use only standard models of stokers and furnaces representative of those widely used.

Jan 1, 1942

By Walter Knox, J. D. Doherty

In 1939, at the request of The Koppers Coal Co., the Koppers Company Research Department established a Stoker Coal Research Laboratory for the purpose of investigating the performance characteristics of stoker coals in typical domestic stoker installations. Fig. I is a photograph of part of the laboratory. The material in this paper deals with the performance of one coal in a single installation. Presented are details of this installation, mechanical characteristics of the stoker, and the results of a series of burning tests in which adjustments were systematically varied. Summary OF Results Mechanical Features I. A large percentage of the total power consumed by a domestic stoker is used to operate the gear case and motor (Fig. 3)- 2' The power consumption per ton goes down with an increase in rate of coal feed (Fig. 4). 3. The automatic air control in this stoker showed advantages Over manual control (Fig. 5). 4. The air velocities through the retort tuyeres were uniform around the retort, but varied between the rows of tuyere rings (Fig. 6). 5. The crushing of coal with this stoker increased the minus 1/8-in. from 11.6 per cent to 19.2 per cent. There was a notice- able segregation of coal as it came from the retort and distribution was not uniform (Fig. 7). 6. Measurement of air entering over the fire showed that with some not uncommon adjustments this air may be a high percentage of the total air supply (Fig. 8 and Table I). Operating Performance ,. The, removal of clinker was shown to give a period of poor performance characterized by a high fuel-bed resistance and a high percentage of carbon monoxide in the flue gases (Fig. 10). 2. The fuel-bed resistance was found to decrease as the fuel bed built up in a normal response to heat demand after a prolonged hold-fire (Fig. II). 3. with intermittent operation the burning rate was almost always less than the feeding rate, The wide variations were found to be greatly dependent upon the length of the On period and the length of the intervals between On periods (Fig. 12). 4. The maintenance of a low furnace draft led to less fly ash, better clinker formation, and higher efficiencies (Fig. 13). 5. The maintenance of a minimum of excess air also resulted in less fly ash, better clinker formation, and higher efficiencies (Fig. 14). Guiding Principles In planning the investigations, the following principles were adopted: Selection of Heating Equipment.—Use only standard models of stokers and furnaces representative of those widely used.

Jan 1, 1942

By R. T. Hukki, A. M. Gaudin

Previous work on the principles of comminution has shown: (I) that the surface produced is proportional to work input (Rittinger law,1a-3); (2) that there is regularity to the weight distribution of the fragments according to size.4-10 In this paper it is brought out that the surface on the fragments is distributed in definite and simple relation to size. The experimental work has involved: I. Construction of an apparatus capable of measuring the surface of mineral fragments independently of the chemical composition of the fragments and independently of the site of the surface; 2. Refining the methods available for measurement of particle size; 3. Designing and building a crusher capable of operating on very small quantities of mineral, if necessary on single crystals, and in such fashion as to minimize handling losses of material. Complete experiments have been carried out with massive quartz and individual quartz crystals. partial experiments, lacking surface measurements, have been made with glass spheres, galena, sphalerite, fluo-rite, salt and coal. Introduction The existence of regularity in weight distribution with size was noted many years ago, especially by A. 0. Gates;' the subject was also studied in detail by one of us.6 In the latter paper it was shown that rocks can be divided into two classes: homogeneous rocks, which show no preferential breaking, and heterogeneous rocks, which show preferential breaking at the grain size of the rock. It was shown, furthermore, that the fragments obtained by rupturing a homogeneous rock occur (in the fine range) in such abundance as to permit the writing of a relatively simple equation to relate the weight retained on a screen to the screen opening. With screens of which the openings are in constant ratio, the weight retained, w, is related to the screen size x by the equation w = Cxm [I] In this equation C and m are experimental constants. A plot of weight retained against particle size is a straight line of slope m if the scales for x and w are both logarithmic. In the fine range, the slopes of the size-distribution graphs plotted according to the method described above vary within a limited range. An inspection of a number of size-distribution data by various experimenters indicates that the slope very seldom exceeds 1.0, but is usually greater than 0.7. More recently, schuhmann8 proposed a method of charting that has so many advantages of its own that it has been used primarily in this work. For fine particles Schuhmann has found that the data are represented by the equation (x)m k [2]

Jan 1, 1947

By A. M. Gaudin, R. T. Hukki

Previous work on the principles of comminution has shown: (I) that the surface produced is proportional to work input (Rittinger law,1a-3); (2) that there is regularity to the weight distribution of the fragments according to size.4-10 In this paper it is brought out that the surface on the fragments is distributed in definite and simple relation to size. The experimental work has involved: I. Construction of an apparatus capable of measuring the surface of mineral fragments independently of the chemical composition of the fragments and independently of the site of the surface; 2. Refining the methods available for measurement of particle size; 3. Designing and building a crusher capable of operating on very small quantities of mineral, if necessary on single crystals, and in such fashion as to minimize handling losses of material. Complete experiments have been carried out with massive quartz and individual quartz crystals. partial experiments, lacking surface measurements, have been made with glass spheres, galena, sphalerite, fluo-rite, salt and coal. Introduction The existence of regularity in weight distribution with size was noted many years ago, especially by A. 0. Gates;' the subject was also studied in detail by one of us.6 In the latter paper it was shown that rocks can be divided into two classes: homogeneous rocks, which show no preferential breaking, and heterogeneous rocks, which show preferential breaking at the grain size of the rock. It was shown, furthermore, that the fragments obtained by rupturing a homogeneous rock occur (in the fine range) in such abundance as to permit the writing of a relatively simple equation to relate the weight retained on a screen to the screen opening. With screens of which the openings are in constant ratio, the weight retained, w, is related to the screen size x by the equation w = Cxm [I] In this equation C and m are experimental constants. A plot of weight retained against particle size is a straight line of slope m if the scales for x and w are both logarithmic. In the fine range, the slopes of the size-distribution graphs plotted according to the method described above vary within a limited range. An inspection of a number of size-distribution data by various experimenters indicates that the slope very seldom exceeds 1.0, but is usually greater than 0.7. More recently, schuhmann8 proposed a method of charting that has so many advantages of its own that it has been used primarily in this work. For fine particles Schuhmann has found that the data are represented by the equation (x)m k [2]

Jan 1, 1947

By Ernest Darby

WHEN bronze is melted in open-flame furnaces a considerable amount of slag is formed during the melting operation. This slag may be incidental to the melting practice or it may be formed intentionally by slag-forming constituents added to the furnace charge to provide a covering for the melted metal. When this slag is removed from the furnace, it contains, in addition to the free metal carried with it by the skimming operation, a considerable amount of metal as oxide either combined or mechanically mixed with it. By the usual crushing and washing methods this metallic oxide generally goes to the tailings pile, so diluted by other foundry waste that its recovery is economically impossible. If this furnace slag is not mixed with other foundry waste, practically all of the metal oxide as well as the free metal may be recovered by some form of smelting operation. The first side of the problem to be considered is the amount of slag produced. It may not be true that the majority of foundries have no record of the amount of slag made, but it is certainly true that a very great many either have no record or pay little attention to it. Usually the collection of furnace slag for one normal week's run will be sufficient to determine the advisability of its being handled in a different manner from the crushing and concentrating generally employed. If a large representative sample of the accumulation is crushed and concentrated and the recovery compared with the results of a careful assay, it will be found that a large percentage of the total metal content was lost in the washing process. The percentage of loss will be greater than the metallic oxide content, as some free metal in extremely fine particles always is carried away in grains of slag. From such a comparison, it may be determined whether or not the slag should be melted, sold or treated as before. An example of an actual test of this kind is given below.

Jan 1, 1928

By J. Philip Neal, Carroll P. Rogers

Feldspars, the most abundant minerals of the igneous rocks, occur in numerous forms and mixtures. The feldspars of commercial significance are found in widely distributed pegmatites as large crystals usually free of iron- bearing impurities and thus suitable for hand-cobbing, and also in larger bodies where the ore contains various types of feldspar inter- mingled with quartz and relatively free of iron- bearing impurities, or at least readily unlocked from impurities present at relatively coarse mesh. The latter type of ore includes some pegmatites, and the ''alaskite" bodies so common in the Spruce Pine district of North Carolina. On the west coast of the U.S. and in Oklahoma there is-r has been--commercial recovery of natural feldspathic sand as a source of alumina in glass and ceramics. "Aplite" in commercial terminology is a light-colored igneous rock with a granitic composition and a fine sugary texture, often banded with readily removable iron impurities. Two areas in Virginia produce aplite for the glass industry: one located in the vicinity of Piney River, Amherst County, and the other near Montpelier in Hanover County. According to Deer et al. (1966) the word feldspar derives from the Swedish feldt or falt (field) and "spath:" chunks of rock appearing in tilled ground overlying granite. Castle and Gillson (1960) indicate a Germanic origin, citing "spat," which is said to refer to any transparent or translucent material which is readily cleavable. The term "spar" has in the past been applied to minerals other than feld- spar, such as barite, calcite, and fluorite. It is correct to call the latter "fluorspar." There is a true barium feldspar, but it is rare and of no present economic importance. What is recognized as feldspar or "spar" by present producers and consumers consists of three silicate minerals which, if pure, would have the formu- las KAISi,O, (microcline or orthoclase), NaAlSi,O, (albite), and CaAl,Si,O, (anorthite). These are almost never found in pure form in nature, but occur together in great abundance in a three-component system. A considerable number of different combined ratios of these three exists, with the exception that isomorphism between potassium spar and calcium spar is very limited. User specifications for feldspar products are largely built around the prevalent naturally occurring ratios. Pas- sages following will refer to "K-spar," "Na- spar," or "Ca-spar." Nomenclature and Descriptive Terms Relating to both the mineralogy and the economics of feldspar are certain terms whose coverage it is useful to specify. Commercially, aplite is a feldspathic rock mined and beneficiated in two locations in Virginia, in which both titanium and feldspar minerals are present-the latter now being the major, if not the only, economic product. Technically, the term aplite when applied to this rock is of questionable accuracy. But, in light of accepted commercial usage, the term will be here used to designate it. Regarding the Virginia rock, Ross (1941) is cited.

Jan 1, 1975

By Ernest Darby

WHEN bronze is melted in open-flame furnaces a considerable amount of slag is formed during the melting operation. This slag may be incidental to the melting practice or it may be formed intentionally by slag-forming constituents added to the furnace charge to provide a covering for the melted metal. When this slag is removed from the furnace, it contains, in addition to the free metal carried with it by the skimming operation, a considerable amount of metal as oxide either combined or mechanically mixed with it. By the usual crushing and washing methods this metallic oxide generally goes to the tailings pile, so diluted by other foundry waste that its recovery is economically impossible. If this furnace slag is not mixed with other foundry waste, practically all of the metal oxide as well as the free metal may be recovered by some form of smelting operation. The first side of the problem to be considered is the amount of slag produced. It may not be true that the majority of foundries have no record of the amount of slag made, but it is certainly true that a very great many either have no record or pay little attention to it. Usually the collection of furnace slag for one normal week's run will be sufficient to determine the advisability of its being handled in a different manner from the crushing and concentrating generally employed. If a large representative sample of the accumulation is crushed and concentrated and the recovery compared with the results of a careful assay, it will be found that a large percentage of the total metal content was lost in the washing process. The percentage of loss will be greater than the metallic oxide content, as some free metal in extremely fine particles always is carried away in grains of slag: From such a comparison, it may be determined whether or not the slag should be melted, sold or treated as before. An example of an actual test of this kind is given below.

Jan 1, 1928

By J. S. Hardison

In the fall of 1912, the appearance of water in the oil of the Nevada Petroleum Co., Coalinga, Cal., made necessary the installation of a dehydrating plant to reduce the water below the 3 per cent. limit prescribed by the agency. Unlike the mining industry, technical literature of the oil industry is limited and extremely unsatisfactory. Until the recent efforts of the Petroleum and Gas Committee of the American Institute of Mining Engineers, no concerted movement has been instituted to secure publication of papers dealing with the problems of the oil business, and because of the additional fact that dehydrating of oil has not been practiced in California to any extent up to comparatively recent time, it was found necessary to experiment in order to determine the most satisfactory plan for this purpose. It is easy to write regarding successful enterprises, but, while not SO pleasant, it is equally desirable to write of failures so that others may be saved the loss incident to such investments. The water in this oil occurs both free and as an emulsion. Free water easily settles out, but the emulsion requires treatment. So far as can be determined, the emulsion consists of globules of water surrounded and enveloped by a film of oil. Starting with this hypothesis the theory has been evolved that in order to break up this emulsion the water must be heated at least to boiling point, when an explosion takes place destroying the globule. Unfortunately, this condition was not recognized in the installation of the first plant, which was planned on the following lines: A heater was arranged so that all the oil from the wells could be heated before reaching the shipping tanks, these being already fitted with coils for steam heating. Tanks were filled with about 4 ft. of water, live steam was turned into the coils and the water brought nearly to the boiling point, when the heated oil was run in. This for a time was partly successful, but the length of time necessary to apply this heat was too great for practical operating.

Jan 1, 1915

[E. O'C. Acker '86 Jinzoc Adachi ‘88 *Truman H. Aldrich '78 Walter H. Aldridge '89 Anson W. Allen ‘87 *John H. Allen '84 William R. Appleby '90 *W. Lawrence Austin '82 *W. SAyres ‘73 David Baker '86 *John Henry Banks .83 Henry C.Banon '90 Alonzo F.. Bardwell '89 Walter A. Barrows ‘90]

Jan 1, 1940

By C. H. Mathewson

EXCEPTING the joint sessions on gases in metals held during the day on Tuesday, the Institute of .Metals opened its activities with a division dinner at the Commodore on Tuesday evening, with Sam Tour, retiring chairman, presiding. W. M. Corse, secretary-treasurer, outlined the financial status of the division and stated the investment fund is becoming large enough so that in the near future it will be possible to pay all expenses of a foreign lecturer once in two years and not rely on an extended speaking tour to cover expenses. W. B. Price reported on his attendance at the Zurich, Switzerland, meeting of the International Society of Testing Materials and gave a graphic account of his visit to the Grand Salon, his meeting with Mussolini, and his audience with the Pope. Buffalo was announced as the place of the fall meeting, to be held during the week of Oct. 3.

Jan 1, 1932

By R. L. Vaughn, A. G. Moon

In 1972 a major renovation of the flotation and regrinding circuits was completed at Kennecott Copper Corporation's Chino Mines Division concentrator at Hurley, New Mexico. At that time a new lime distribution and control system was placed in operation. Lime is used to depress pyrite, and flotation at a high pH or alkalinity is required for optimum recovery of principal copper minerals, chalcopyrite and chalcocite with good rejection of the pyrite. Regrinding is also required to liberate chalcocite, much of which occurs as intimately associated chalcocite-pyrite middling grains. Accurate control of the lime addition is necessary to obtain both optimum metallurgy and minimum lime usage. At low pH levels, flotation cell control is difficult due to excess froth, but excessively high pH levels can go unnoticed because froth characteristics do not change from those experienced at the desired pH. With manual control of the lime addition, the desired pH can be exceeded for periods of up to two hours. Test Work Plant test work on a lime control system using conventional pH probes gave disappointing results with high electrode maintenance and poor control characteristics. For example, the rougher flotation cells are operated at a pH between 11.4 and 11.6 and the probes required cleaning at least once a day, and occasionally every shift. These high maintenance requirements were considered unreasonable for a plant-wide system using six or more control loops.

Jan 1, 1979

By H. R. Page, A. E. Jr Lee

A. E. LEE, JR. (author)—In addition to the paper we should like to make a few remarks. First, the seriousness of bending of the clay retort cannot be overemphasized. Not only did bending limit the length of the vessel, but it permitted corrosive fluid slags to concentrate in the belly of the retort and thus increased the possibility of boring. Bending also practically prohibited the use of cleanout machines and increased the danger of excessive fume losses by allowing furnace gases to escape around the top of the retort mouth. It will be noted that the Blackwell pottery practice places considerable emphasis on careful drying of retorts. Perhaps it is another smelter superstition, but traditionally the aging of retorts beyond a reasonable drying period has been considered important to long life. We are unable to give a good explanation of why aging should improve retorts. During the past fall in an attempt to prove a theory that the combined water in the clay decreased during aging, about 50 samples of wet retort collars were placed in a dry room when it was closed. Loss of weight, combined water and free moisture were measured each week over the 100 day drying period. The affinity of the dried retort for moisture makes difficult the accurate measurement of free water; however, there was a clearly defined free moisture change from about 12.5 pct when freshly made to about 2 pet at the end of 18 days, then a gradual loss to about 0.25 pet at 70 days and an apparent equilibrium near this figure through the 100th day. There was no appreciable change in the combined water. The change in weight of the average sample was, however, rather unusual, there being consistently a loss of weight throughout the entire 100 days. From the 63rd day on this leveled off to about 0.03 pet per week. These were small samples and certainly should have dried

Jan 1, 1950

ALABAMA Altoona.-Cain, J. America.-Foreman, J. T. Anniston.-Cowie, L. K. Foster, R. N. Rogers, R. F. White, H. E. Ashland.-Sturdevant, J. C. Bessemer. Ball, E. M. McKenzie. W. C., Jr. Morgan, C. Birmingham.-Abbott, C. E. Adler, J. E. Aldrich, T. H. Aldrich, T. H., Jr. Ashton, J. J. Burgin, G. E.

Jan 1, 1923