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By Fred C. Bond and

This paper is a continuation of two earlier papers,12 and presents new data on the grindability of various ores and other materials—the results of several years of intermittent research work on the nature of grinding and interpretation of test results. It is divided into five parts: 1. The tabulated results of several hundred grindability tests on a wide variety of ores. 2. A method of correcting screen analyses and calculating from them the surface and other properties. 3. A description of a new method of measuring the "surface energy" of an ore by pendulum impact tests. 4. Test results. 5. A general discussion of the test results and their application to grinding problems. 1. GRindability of Various Ores This part is a continuation of the paper listed as reference No. 1. The method of testing is described in detail in that paper. The sample to be tested is crushed to all through 6 mesh in rolls set at 1/16-in. opening in closed circuit with a screen. Of this sample 700 C.C. is ground dry in a laboratory ball mill under constant conditions of mill speed and ball charge in closed circuit with a screen of the desired size at 250 per cent circulating load. The mill is cylindrical, 12 by 12 in., and rotates at 70 r.p.m. It contains a charge of 285 iron balls weighing 20,125 grams and ranging from 1½ to ¾ in. in diameter. Each test is continued for a sufficient number of periods to establish equilibrium at 250 per cent circulating load, or until the composition of the circulating load is constant for successive periods. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the results are expressed as the net production in grams of screen undersize per revolution of the mill. This is designated as the

Jan 1, 1939

By Arthur F. Taggart

This paper proposes two units (believed to be new) for designating, respectively, capacity and efficiency for primary and intermediate crushers. Capacity Operators know that the tonnage of rock that can be put through a particular crusher in a given time depends primarily upon the crusher setting (open setting for jaw and primary gyratory types; closed setting for cones and rolls), and secondarily on the lithological character of the feed, its maximum size, the amount of fines that it carries, its moisture content, and the method of feeding. The proposed capacity formula states a relationship between all of these factors in one empirical equation, as follows: Tr = TR80K'K" [I] in which Tr, called "reduction tons per hour" is the new capacity unit; T = tons of feed per hour requiring crushing; R80 = "80-per cent reduction ratio"; and K, K', and K" are constants that approximate quantitatively lithological character, moisture content and method of feeding, respectively. T is the hourly weight of material in the feed that is coarser than the coarsest discharge of the crusher. (For marked departures from the normal specific-gravity range—2.5 to 2.7—T should be adjusted to corresponding volumes.) In the crushers under consideration, discharge from the crushing zone is limited by the minimum dimension of a particle. Passage of the particle through a square-mesh screen is limited by its intermediate dimension. Sheppardl has shown that the ratio of maximum: intermediate: minimum average dimensions of broken rock fragments may be taken as about 1.5:=:0,6 for nonslabby rocks such as are normally crushed, while for slabby rocks, such as thin-bedded sedimentaries, foliated schists and gneisses, and the blocky metamorphics, the ratio is of the average order I.3:I:o.3, with the possibility of an even smaller relative value for the thickness. Gaudin2 finds the corresponding ratio for —I50 + 200-mesh pyrite to be I.4: I:0.6. The hourly tonnage of crusher-plant feed, therefore, should be diminished by the amount thereof that would pass a square-mesh screen of which the aperture is S times the crusher setting, where S is a shape factor with the value 1.7 for rocks that break exceptionally cubic; 2 for ores and rocks that exhibit no appreciabie tendency to slabbiness; and 2.5 to 3, or, in exceptional cases, even larger, when the tendency to slabbiness is marked. R80 is the quotient of the theoretical square-mesh aperture that would pass 80 per cent of the feed, divided by the aperture that would pass 80 per cent of the product. It is best determined from screen tests. If, as is usual with primary crushers, a sizing test of feed is not available, the

Jan 1, 1943

By Arthur F. Taggart

This paper proposes two units (believed to be new) for designating, respectively, capacity and efficiency for primary and intermediate crushers. Capacity Operators know that the tonnage of rock that can be put through a particular crusher in a given time depends primarily upon the crusher setting (open setting for jaw and primary gyratory types; closed setting for cones and rolls), and secondarily on the lithological character of the feed, its maximum size, the amount of fines that it carries, its moisture content, and the method of feeding. The proposed capacity formula states a relationship between all of these factors in one empirical equation, as follows: Tr = TR80K'K" [I] in which Tr, called "reduction tons per hour" is the new capacity unit; T = tons of feed per hour requiring crushing; R80 = "80-per cent reduction ratio"; and K, K', and K" are constants that approximate quantitatively lithological character, moisture content and method of feeding, respectively. T is the hourly weight of material in the feed that is coarser than the coarsest discharge of the crusher. (For marked departures from the normal specific-gravity range—2.5 to 2.7—T should be adjusted to corresponding volumes.) In the crushers under consideration, discharge from the crushing zone is limited by the minimum dimension of a particle. Passage of the particle through a square-mesh screen is limited by its intermediate dimension. Sheppardl has shown that the ratio of maximum: intermediate: minimum average dimensions of broken rock fragments may be taken as about 1.5:=:0,6 for nonslabby rocks such as are normally crushed, while for slabby rocks, such as thin-bedded sedimentaries, foliated schists and gneisses, and the blocky metamorphics, the ratio is of the average order I.3:I:o.3, with the possibility of an even smaller relative value for the thickness. Gaudin2 finds the corresponding ratio for —I50 + 200-mesh pyrite to be I.4: I:0.6. The hourly tonnage of crusher-plant feed, therefore, should be diminished by the amount thereof that would pass a square-mesh screen of which the aperture is S times the crusher setting, where S is a shape factor with the value 1.7 for rocks that break exceptionally cubic; 2 for ores and rocks that exhibit no appreciabie tendency to slabbiness; and 2.5 to 3, or, in exceptional cases, even larger, when the tendency to slabbiness is marked. R80 is the quotient of the theoretical square-mesh aperture that would pass 80 per cent of the feed, divided by the aperture that would pass 80 per cent of the product. It is best determined from screen tests. If, as is usual with primary crushers, a sizing test of feed is not available, the

Jan 1, 1943

By F. Cadena

HIWASSEE Dam, now under construction by the Tennessee Valley Authority on the Hiwassee River, a tributary of the Tennessee River, will require aggregate for approximately 800,000 cu. yd. of concrete. Near the damsite deposits of natural sand or gravel are not available and graywacke, a rock that is found in abundance, suggested itself as a likely source of manufactured aggregate. Graywacke is a metamorphosed sandstone composed principally of a mixture of quartz (65 per cent), mica (25 per cent), and feldspar (10 per cent). Other minerals are pres-ent in negligible amounts. The rock varies in different parts of the quarry from very fine to a coarse grain containing particles of about 1/4 in. maximum size, although the bulk of the deposit is medium grain. Exten-sive tests showed that graywacke would be suitable as a source of both coarse and fine aggregates. The aggregate plant available for use at Hiwassee Dam consisted of the coarse-aggregate and sand plant used on the Norris project. The coarse-aggregate plant had also been used at Chickamauga Dam. Although most of this plant equipment is now in use at Hiwassee, the present flowsheet is not only different from that at Norris and Chicka-mauga but is also unique as compared to the flowsheet of the average crushed-stone aggregate plant. It is hoped that this paper will show the importance of proper arrange-ment of equipment if the optimum in flexibility and capacity is to be attained.

Jan 1, 1939

By H. M. Otte, A. L. Esquivel

A new leasl-squares method is Presented for determining lattice parameters of hexagonal or tetragonul structures. The method is adapted for use on electronic computers and involves a reiterative procedure. The correction factor employed raries linearly with the lattice parameter, a (determined from the Brag, angle). In contrast, Cohen's method and recent modifications of it use a correction factor that varies incersely as the squure of the lattice paramneter, a. While the recent modifications attempt to improve the precision of the extrapolated lattice parameter, a, (or-cu). by stressing the importance of the weighting factor. the present approach emphasizes the need for choosing the correct extrapolation function. A comparison between the present method (the Linear method) and Cohen's method indicates that the Linear method may be more appropriale in certain cases. through a priori no critertion appears to he available for making a chorce between the methods. The size of hexagonal and tetragonal crystal lattices is determined by two parameters, a, and c,. Although in principle the two lattice parameters can be determined independently from reflections for which h - k = 0 and 1 = 0, respectively, in practice this may be inconvenient (because of the angular positions at which these reflections may occur) or not easily possible (because of low intensity). Furthermore, if a high precision or accuracy is required, the limited number of reflections of this type available, particularly in the high-angle region, is not sufficient for the necessary corrections (mainly due to absorption) to be determined with accuracy. Several methods1-7,11-13 have been proposed employing all reflections, to obtain the optimum values, a. and co, either by a trial and error procedure or a least-squares fit. Of the latter method, cohen's5 is the best known one since it provides explicit expressions for the optimum a. and c,. However, Cohen's method is only strictly valid if an extrapolation function is used that varies linearly with l/a2 or lie2 (see Section 3), a requirement that does not appear to be generally appreciated. On the other hand, all the better known and more widely used A recent trial and error method was proposed by Massalski and King8,7 who computed extensive auxiliary tables of axial ratios vs the functions A = [(4/3)(h2 + hk +k2) + l2+(a/c)2] and C = A(c/a)2 used in computing a and c values from the observed Bragg angles. These values of a and c were then plotted against a function which permitted linear extrapolation. As a criteria; for the "optimum" values, Massalski and King rely upon a visual fitting of the line through points representing reflections of low 1-index points to compute the extrapolated value of ao and high 1-index reflections to obtain CO. The successive computations and graphical plotting required to reach the "optimum" value are quite lengthy and tedious even on a desk calculator and no quantitative assurance is obtained of having in fact selected the optimum value.* If the method of Massalski and King is used on an electronic computer, then their published tables become redundant and a least-squares fit becomes a natural selection for the choice of optimum values. Such an approach will be called the Linear method. For work now in progress on the effect of certain physical variables on the lattice parameters of hexagonal crystals, it has become essential to determine the confidence limits of small changes in the lattice parameters. Since extrapolation functions that varied linearly with a and c actually also appeared to vary linearly with l/a2 or l/c2 when tested against published as well as our data, a comparison of Cohen's method and the Linear method was considered desirable (Sections 3 and 4). For the latter method an electronic computer was required since a reiterative procedure to obtain the optimum ao and co values had to be employed. The purpose of this paper is to describe the principles of the Linear method, illustrate its application, and compare it with Cohen's method. 1) THE LINEAR METHOD The standard practice in obtaining the lattice parameters in the Debye-Scherrer method is to

Jan 1, 1965

By Eric Kula, Åke Josefsson

L. J. Dijkstra and R. Sladek, (Ontario Research Foundation, Toronto, and Institute for the Study of Metals, Chicago, respectively)—This interesting paper confirms some results obtained some years ago in collaboration with Fast: namely the broadening of the nitrogen peak by the presence of small amounts of manganese. Later on a more extensive study of this effect was made at the Institute for the Study of Metals at Chicago. As the work has not yet been submitted for publication we wish to mention here briefly some of the results. The effect of about 0.5 atomic pct manganese, chromium, molybdenum, or vanadium on the nitrogen-peak in iron was examined. The broadening found previously in the case of manganese was even found to be more pronounced for chromium. In the case of molybdenum and vanadium a double peak was observed which seemed to justify the idea that in all cases mainly two relaxation times were involved: namely, the normal relaxation time as found in pure iron and a second relaxation time related to the existence of abnormal interstitial lattice sites introduced by the foreign metallic element. The analysis showed that for the normal nitrogen-peak as a reference at 22°C the abnormal peak was found at about 32°, 47°, 75°, and 87°C in the case of manganese, chromium, molybdenum, and vanadium, respestively. In the manganese alloy the free energy difference for a nitrogen-atom at an abnormal and a normal lattice site was estimated at 2800 cal per mol. This means that for temperatures above 100°C the solid solubility and overall diffusion rate of nitrogen is not much affected by additions of 0.5 pct Mn. We believe that the difference in rate of precipitation of nitrogen in pure iron and the manganese alloy is mainly a question of the number of potential nuclei available in the different steels. In studying the effect of these alloying elements on precipitation of nitrogen, a distinction must be made between the para reaction and ortho reaction, using terminology introduced by Hultgren.7 The ortho reaction in which the alloying metallic element participates in the diffusion under formation of nitrides occurs at high temperatures such as 600°C. The para reaction in which the alloying metallic element does not diffuse and in which the nitride inherits its alloy content from the matrix occurs at lower temperatures. To prevent the ortho reaction the alloys have to be quenched rapidly from the phase. It is not surprising that the authors find that the height of the peak indicates a lower nitrogen content than chemical analysis. In the first place the peak is not a single relaxation peak and secondly, the steels were not quenched from the phase after the nitrogen was introduced at 580 °C.

Jan 1, 1953

By J. Lohrenz, R. J. Francis, G. C. Clark

Compositional material balance methods have been extended to solution gas-gas cap-water drive reservoirs. Partial or complete pressure maintenance as well as reservoir repressuring by gas and/or water injection may be considered. Increrments of reservoir withdrawals and injections equivalent to a given time increment are used as the basic iterative step to allow the extension. An example field study shows how the reservoir volumes. extent of the water drive, and relative permeability ratio function may be estimated. Knowing these, the methods can be used to predict reservoir behavior. INTRODUCTION The compositional material balance is similar to conventional material balances in many aspects. The assumption of a pressure-homogenous reservoir with no transients is common to all conventional material balances1-1 although numerical methods are different. The compositional material balance also contains this assumption. Inclusion of phase equilibria directly in reservoir calculations is the key difference and improvement in the compositional material balance. A continuous account of the compositions of the reservoir fluids and effluents is determined. Calculations of fluid phases and properties can be based on these compositions. Excepting fluid phases and properties determined in this manner, the compositional material balance uses procedures exactly analogous to conventional methods. Brinkman and Weinaug5 and Jacoby and Berry6 have given compositional material balances for simple solution gas drive reservoirs; subsequently Jacoby and Berry' extended their procedure to include pressure maintenance by gas injection at constant reservoir pressure. This paper extends compositional material balance techniques to a more general combination drive reservoir with various types of pressure maintenance operations. The procedure for this extended compositional material balance and an example of its field application are presented. PROCEDURE The numerical calculations of this compositional mate- rial balance are performed on a high-speed digital computer. The most comprehensive case the present compositional material balance program can handle is a solution gas-gas cap-water drive reservoir with (1) water production, (2) water influx, (3) separate or commingled production from the oil and gas zones, (4) any surface processing arrangement including LPG and gasoline production, and (5) gas injection into the gas and/or oil zones." This description of the procedure applies to this most comprehensive case. Appropriate deletions in the procedure can be made for more simple mechanisms. ASSUMPTIONS In common with conventional material balances, the compositional material balance makes the following assumptions: (1) the total reservoir is at a uniform pressure at a given time; (2) the volumetric ratio of gas to liquid withdrawal at reservoir conditions is directly proportional to their mobilities; and (3) instead of assuming a pressure-dependent total gas-oil equilibrium, the cornpbsitional material balance makes a sianificantly different and more rigorous assumption as follows. Thermodynamic equilibrium exists for each component in the liquid-gas mixtures in the reservoir and in surface processing. These three assumptions are fundamental in the compositional material balance procedure. Other assumptions involve arbitrary simplifications of the reservoir mechanisms. This compositional material balance uses the following simplifying assumptions: (1) no bulk mixing occurs between the gas and oil zone; (2) total mixing occurs between any injected gas and the contents of the injection zone; (3) no in situ hydrocarbon liquid is produced from the gas zone; (4) no reservoir gas is retained behind any net bottom-water influx; (5) the residual saturation of hydrocarbon liquid at reservoir conditions behind any net bottom-water influx is constant; and (6) fluids retained behind any net bottom-water influx exert no further influence on the reservoir. These six assumptions are not fundamental. Modifications of the present procedure can be made for different assumptions of the reservoir mechanisms. THE TIME INCREMENT Previous compositional material balances have used a pressure increment for simple solution gas drive reservoirs"" and an increment of injected gas for pressure maintenance.' These increments are suitable for these simple cases. They are, however, inadequate for reservoirs with time dependent drive mechanisms including water drive. A time increment removes this inadequacy and is

By G. M. T. Marshall

Plans to bring the Geco Mines property in Manitouwadge, Ont., into production began in 1954. At that time over 12 million tons of copper-zinc ore were indicated in a zone 2700 ft long. A concentrating plant was built to handle 3300 tpd of ore and a production shaft was sunk about 1500 ft. MINE LAYOUT Vertical attitude, width, rake, and competent country rock of a 4-million ton orebody west of the main diabase dike suggested a system of blasthole stopes, undercut with scram drifts connected by a transfer raise system to a crusher on the 1250-ft level and this system was employed. A 42-in. conveyor on this level transfers the ore about 1200 ft to a loading pocket at the No. 1 shaft. The system has a number of advantages: 1) Cost of tramming is eliminated. 2) Rate of extraction from any stope is controlled only by the scraping capacity of the scram. 3) Rate of extraction from the stoping area can be very rapid. Fast extraction makes it possible to cut mining costs for the following reasons: The number of stopes actively producing is kept to an economical minimum; the operating area is concentrated-manpower can be used more effectively; and active life of the stoping block is short, making ground support problems simpler and less costly.

Jan 8, 1959

By Rudolph Emmel

The mines operated by the South American Development Co. are located in the Zaruma mining district of southwestern Ecuador. They are near the old mining town of Zaruma, which is the only important city in the canton of the same name. The district is situated in the upper end of a valley lying between two spurs of the Cordillera, one of which may be considered the Coast Range and the other an intermediate range. The mining camp proper, known as Portovelo, and the plant are 2.4 km. south of Zaruma, on the north bank of the Amarillo River, a tributary of the Tumbes River, which flows southwestward through Peru to the Pacific. Portovelo is difficult of access; it is reached from Guayaquil by means of river steamers and muleback. 'Embarking at Guayaquil, the route is down the Guayas River, across the Jambeli Canal, and up the Santa Rosa River to Santa Rosa, a distance of 177 km. From Santa Rosa, the road or trail follows the Santa Rosa River, then across the summit of the Coast Range into Portovelo, a distance of 74 km.—a two-day trip on muleback. The mining property of the South American Development Co. comprises 297 lode claims and 142 placer claims. The combined lode claims cover an area, roughly rectangular in shape, 9000 m. long from north to south by 4000 m. wide; the northern edge of the rectangle is about 3000 m. north of the central plaza of Zaruma. Mineral lands in Ecuador are held by virtue of an annual tax per claim paid to the central government, no surface rights being included. HiSTOry The Zaruma mine has been worked by white men since 1549, probably before that it was worked by Indians. The followers of Pizarro, led by traces of gold in the sands of the Tumbes River, followed that stream to its source, worked the gold veins that they found and established the city of Zaruma. Among the mines worked by them were th Sesmo, Leonora, Viscaya, and Mina Grande. Only the soft ores at the outcrop were

Jan 1, 1925

By I. N. Zavarine

The observations presented in this paper were recorded during a study of the spheroidizing process. Spheroidization of cementite in steel is either brought about to develop a set of desirable mechanical properties or is an incidental result of mechanical and heat-treating manipulations. Physical properties of steel are materially influenced by the state of the cementite. Conversion of cementite from the lamellar to the spheroidal state is usually accompanied by a decrease of hardness and of strength and by an increase in ductility. The spheroidized structure is usually considered as an ultimate step in the process of the softening of steel. The unorthodox hardening of a 3.5 per cent Ni steel (SAE No. 2330) after a heat-treatment that previously proved entirely suitable for a complete spheroidization of cementite in carbon steels led to the observations that constitute the subject of the present paper. Experimental Procedure The steels used for the experiments described were bought on the market. All were in the form of 3/4-in. round bars. Their chemical compositions are given in Table 1. The spheroidization of cementite in steels can be obtained in several ways. The method used in the present work is, perhaps, the least known. It consists of cyclic heating and cooling of steel, the temperature being

Jan 1, 1939

By I. N. Zavarine

The observations presented in this paper were recorded during a study of the spheroidizing process. Spheroidization of cementite in steel is either brought about to develop a set of desirable mechanical properties or is an incidental result of mechanical and heat-treating manipulations. Physical properties of steel are materially influenced by the state of the cementite. Conversion of cementite from the lamellar to the spheroidal state is usually accompanied by a decrease of hardness and of strength and by an increase in ductility. The spheroidized structure is usually considered as an ultimate step in the process of the softening of steel. The unorthodox hardening of a 3.5 per cent Ni steel (SAE No. 2330) after a heat-treatment that previously proved entirely suitable for a complete spheroidization of cementite in carbon steels led to the observations that constitute the subject of the present paper. Experimental Procedure The steels used for the experiments described were bought on the market. All were in the form of 3/4-in. round bars. Their chemical compositions are given in Table 1. The spheroidization of cementite in steels can be obtained in several ways. The method used in the present work is, perhaps, the least known. It consists of cyclic heating and cooling of steel, the temperature being

Jan 1, 1939

By Herbert H. Uhlig

Although the mechanical properties of high-carbon manganese-iron alloys, particularly the Hadfield manganese steels, have been established, the literature discloses discrepancies in the reported properties for the low-carbon alloys. Hadfieldla summarized the existing data up to 1927, including the investigations of Guillet, Arnold, Strauss, Burgess and Aston, and reported his own values for alloys containing for the most part from 0.06 to 0.I5 per cent carbon. He noted: From these comparisons it wil! be seen that while there is a broad and general agreement as to the character of the alloys at the different manganese percentages, there are disparities in individual cases, especially as regards the degree of toughness or brittleness displayed by an alloy with certain percentages and with particular types of treatment. These differences are not always readily explainable.lb His data showed that alloys containing 4 to 10 per cent manganese water-quenched from Iooo°C. were brittle and hard with maximum brittleness at about 7 per cent. Slight improvement in ductility was obtained with compositions from 10 to 17 per cent manganese. Alloys containing up to 17 per cent manganese slowly cooled rather than quenched from 87o°C., he found, were slightly improved in ductility with significant improvement in alloys analyzing 4.1 and 4.8 per cent manganese. Recently Walters, Kramer and Loring2 reported mechanical properties for alloys made up from electrolytic manganese and ingot iron and containing from 0.02 to 0.04 per cent carbon. They showed that a tempering treatment at 54o°C. (Iooo°F.) for one hour significantly improved the ductility of compositions ranging from 6.7 to &apos; 20.7 per cent manganese. The differences of these compared with Hadfield&apos;s results they considered to originate in this tempering treatment plus the fact that Hadfield&apos;s alloys contained higher carbon and phosphorus (0.06 to 0.I5 per cent C; 0.06 to 0.07 per cent P compared with 0.02 to 0.04 per cent C, <o.ooI per cent P). In this paper it is shown that of the low-carbon alloys (0.02 to 0.03 per cent) containing from 3.3 to 21.5 per cent manganese moderately or rapidly quenched from IOOO°C. some are brittle or ductile, depending on presence or absence of hydrogen. Hydrogen is readily absorbed at high temperatures, and much of it is retained when the alloys are cooled to room temperature. The presence of hydrogen in the range of composition from 10 to 18 per cent manganese is a major factor in determining relative ductility. At 22 per cent manganese, and also at compositions below 10 per cent manganese, factors other than or in addition to hydrogen play a role. These are discussed. Also data are presented to show that hydrogen absorbed by the high-carbon alloys of Hadfield composition is without effect on their mechanical properties.

Jan 1, 1944

By Herbert H. Uhlig

Although the mechanical properties of high-carbon manganese-iron alloys, particularly the Hadfield manganese steels, have been established, the literature discloses discrepancies in the reported properties for the low-carbon alloys. Hadfieldla summarized the existing data up to 1927, including the investigations of Guillet, Arnold, Strauss, Burgess and Aston, and reported his own values for alloys containing for the most part from 0.06 to 0.I5 per cent carbon. He noted: From these comparisons it wil! be seen that while there is a broad and general agreement as to the character of the alloys at the different manganese percentages, there are disparities in individual cases, especially as regards the degree of toughness or brittleness displayed by an alloy with certain percentages and with particular types of treatment. These differences are not always readily explainable.lb His data showed that alloys containing 4 to 10 per cent manganese water-quenched from Iooo°C. were brittle and hard with maximum brittleness at about 7 per cent. Slight improvement in ductility was obtained with compositions from 10 to 17 per cent manganese. Alloys containing up to 17 per cent manganese slowly cooled rather than quenched from 87o°C., he found, were slightly improved in ductility with significant improvement in alloys analyzing 4.1 and 4.8 per cent manganese. Recently Walters, Kramer and Loring2 reported mechanical properties for alloys made up from electrolytic manganese and ingot iron and containing from 0.02 to 0.04 per cent carbon. They showed that a tempering treatment at 54o°C. (Iooo°F.) for one hour significantly improved the ductility of compositions ranging from 6.7 to &apos; 20.7 per cent manganese. The differences of these compared with Hadfield&apos;s results they considered to originate in this tempering treatment plus the fact that Hadfield&apos;s alloys contained higher carbon and phosphorus (0.06 to 0.I5 per cent C; 0.06 to 0.07 per cent P compared with 0.02 to 0.04 per cent C, <o.ooI per cent P). In this paper it is shown that of the low-carbon alloys (0.02 to 0.03 per cent) containing from 3.3 to 21.5 per cent manganese moderately or rapidly quenched from IOOO°C. some are brittle or ductile, depending on presence or absence of hydrogen. Hydrogen is readily absorbed at high temperatures, and much of it is retained when the alloys are cooled to room temperature. The presence of hydrogen in the range of composition from 10 to 18 per cent manganese is a major factor in determining relative ductility. At 22 per cent manganese, and also at compositions below 10 per cent manganese, factors other than or in addition to hydrogen play a role. These are discussed. Also data are presented to show that hydrogen absorbed by the high-carbon alloys of Hadfield composition is without effect on their mechanical properties.

Jan 1, 1944

By W. A. Jr. Slippy, E. P. Dahlberg, R. B. Reed-Hill

A large room-temperature mechanical-hysteresis effect under cyclic tensile loading was observed in zivconium specimens prestrained at 77°K so as to form large numbers of (1121) twins. The observed hysteresis loss was a maximum when the prestrain was just under 1 pct It also depended strongly on the maximum applied cyclic stress, but only moderately on loading rate and temperature (-72°to 100°C). The phenotnena have also been studied in terms of the elastic aftereffect where the data has been found to cotlforln closely to the functional relationship between strain and time, c = -RT/a In tank ft. t to)/Zr, for strains cts large as about 10-4 This eyication can be derived directly from the strain-rate equation There may be important commercial implications to the present discovery since it appears that within rather wide limits it may be possible to obtain any desired magnitude of damping capacity in zirconium melal. The damping propevties may also be given directional characteristics. It is shown that the observed anelaslic phenomena can he explained on the assumiplion that they are the resuilt of stress-induced twin-boundary movements in which the average twin increases its thickness by only a few percent. In polycrystalline zirconium, a grain is least favorably oriented for slip when its basal plane, containing the (1130) slip directions, lies nearly perpendicular to the principle normal stress. Mechanical twinning is favored by this orientation and under simple tensile deformation at room temperature the primary twinning mode is {1012).&apos; The (1012) twinning shear is small (0.17) so that, during deformation, twin growth does not greatly enhance ductility. At 77°K, on the other hand, the predominant twinning mode shifts to {1121), and many thin (1121) twins can form at small strains.&apos; Twins thus nucleated can grow readily during subsequent room-temperature deformation and permit easy deformation in unfavorably oriented grains. In addition, the (1131) twinning shear is large: 0.63. These facts have lead to a process&apos; for greatly improving the room-temperature ductility of zirconium when it contains a sizable fraction of grains with basal planes nearly normal to the stress axis. Optimum results are obtained by slightly less than 1 pct prestrain. The effect is large so that a 50 pct elongation increase is readily attained. A large room-temperature mechanical-hysteresis effect has also been observed in zirconium specimens prestrained at 77°K. The present paper is concerned with some aspects of this effect which the experimental observations strongly indicate is primarily the result of strain-induced {1121) twin-boundary movements. This clearly suggests that mechanical twins formed by plastic deformation can cause anelastic effects similar to those previously observed2 for annealing twins and twins associated with phase transformations. EXPERIMENTAL PROCEDURES All specimens were prepared from arc-melted sponge zirconium plate stock as previously described,3 which has a preferred orientation with basal planes aligned parallel to and uniformly distributed around the rolling direction. Both lonti-tudinal and transverse tensile specimens, with axes parallel and transverse, respectively, to the plate rolling direction, were cut from this plate. In a longitudinal specimen most grains are favorably oriented for slip, while in a transverse specimen a large fraction are unfavorably oriented. Prestraining at 77°K was performed in two ways. In one, annealed zirconium plate stock was cooled in liquid nitrogen and then rolled in the transverse plate direction to strains varying from 0.5 to 12 pct. Cylindrical 1/8-in.-diameter by 1-1/4-in.-gage-length tensile specimens were machined from this material. In the other, previously machined tensile specimens were prestrained in tension at 77°K between 0.2 and 3.8 pct. Type FA 25-12 SR4 strain gages were cemented to all specimen gage sections. All tests were performed on an Instron testing machine of 10,000-lb capacity. EXPERIMENTAL RESULTS Fig. 1 shows a typical two-cycle room-temperature stress-strain diagram for a specimen prestrained 0.65 pct by rolling at 77°K. The maximum stress in both cycles was 14,500 psi. Note that in the first cycle the loading and unloading curves are unsymmetrical so that a residual strain, er, remains after unloading. In the second cycle the stress-strain curves form a nearly symmetrical hystere-

Jan 1, 1965

By Robert Ammon

The purpose of this paper is to present a brief sketch of the development of this hindered-settling classifier, but primarily to show the actual results obtained in practice with the classifier working on the Butte copper ores at the Boston & Montana Concentrator at Great Falls and the Washoe Concentrator at Anaconda. The writer has endeavored to be as brief as possible, giving only the essential figures showing the efficiency of the classifier and leaving considerable interesting information to be derived by the reader from the statistics given. The paper deals with the practical work of the classifier under the following heads: 1. Desliming 4.0-mm. primary feed, overflow from 0.07 to 0.0 mm. 2. Desliming 1.25-mm. secondary feed, overflow from 0.07 to 0.0 mm. 3. Desliming Huntington Mill discharge, overflow from 0.07 to 0.0 mm. 4. Classifying 2.5-mm. deslimed secondary feed, overflow from 1.00 to 0.07 mm. 5. Classifying 2.5-mm. deslimed primary feed, overflow from 0.75 to 0.07 mm. 6. Classifying 4.0-mm. deslimed primary feed, overflow from 0.75 to 0.07 mm. 7. Classifying middling-tailing product from roughing tables, overflow coarse tailing. 8. Classifying 2.5mm. trommel undersize, not deslimed, overflow from 0.75 to 0.0 mm. 9. Remarks on the inner feed cone. In the following pages all overflows are given in quartz dimensions; i. e., by a 0.75-mm. overflow is meant an overflow whose maximum particle is a 0.75-mm. quartz grain. By the term "slime" is meant all material which will pass through a 200-mesh screen whose average opening is 0.07 mm. The development of the classifier was commenced in the fall of 1910,

Jan 1, 1914

By R. S. Rickard, B. Ball

INTRODUCTION Pyrite responds to the flotation reagents normally used in sulfide flotation systems. For example, it is well known that xanthates are good collectors for pyrite, and cyanide, sulfide and hydroxides are depressants. The particular response of pyrite to these and other reagents is used to advantage in several areas. By careful control of depressant and collector additions, pyrite can be separated from other sulfides by flotation. This is of obvious importance in base metal sulfide flotation where pyrite depression is often required to produce a marketable grade concentrate. Flotation of pyrite has long been employed in gold ore processing, and it appears that pyrite flotation from coal may become important in the future (1). It is probably a direct consequence of this relative ease of separation, plus the low economic value of pyrite, that detailed investigations on the chemistry of its flotation and depression are relatively limited. Gaudin (2, 3, 4) and others, in the mid-1950&apos;s published data on the flotation of pyrite, but detailed or consistent mechanisms were lacking. In 1955, Sutherland and Wark&apos;s book (5) was published, in which a large , amount of empirical information, especially earlier work of Wark and Cox, was presented. More recently, in 1968, significant contributions have been made by Fuerstenau et al. (6, 7) and Majima and Takeda (8). These papers introduced electrochemical explanations of the effects observed with xanthate and cyanide ions.

Jan 1, 1976

By R. B. Burt, James A. Barr, I. S. Tillotson

THE pumping of solids in water suspension is an important part of many metallurgical and mining operations. In most cases, it is still in the rule of thumb category for which no universal formula has been developed, and much research is needed. Because of the limited and incomplete data available, this article may be classed as an experience paper, which is presented with the hope that some contribution will be made toward the development of the so-called universal formula. This formula, if and when developed, may be evolved from several factors, many of which are not now available for general application. The designing engineer is interested in obtaining accurate forecasts on: 1-the minimum velocities needed to prevent choke-ups in the pipeline, which in turn dictates pipe sizes, 2-power required for pumping, 3-pump selection. The basic factors for a given problem will include: 1-weight per unit of time of solids to be handled, 2-specific gravity of solids, for calculation of volume, friction and power, 3-screen analysis of solids with the colloidal acting, i.e., the slime fraction, a very important factor, 4-shape of particle or some means of determining a friction constant, 5-effects of percentage of solids, 6-development of a viscosity factor to be used in the overall calculations, 7-calculation of the lower limits of pipeline velocities permissible, 8-calculation of total head, pump horsepower, and 9-setting up of pump specifications. In certain limited cases horsepower and total heads and minimum velocities may be computed and a suitable pump selected from basic data, but in many cases, as in mining of Florida pebble, phosphate, experience rather than a hydraulic formula still should be used as a basis of selection. Pumping Florida Pebble Matrix Pumping at the Noralyn mine of International Minerals and Chemical Corp. will be used as an example. Other areas will vary as to the characteristics of the matrix, especially the slime content. A typical screen analysis of this matrix is: +14 mesh, pebble size,* 2.1 pct; -14 +35 mesh, 11.4 pct; -35 +150 mesh, 60.5 pct; -150 mesh, 25.0; total, 100 pct; moisture in bank, 20.0 pct; weight per cu ft in bank, 120 lb. The -150 mesh fraction may increase to as much as 35 pct in adjacent areas. When thoroughly elutriated, the matrix has a relatively slow settling rate, which is an important factor in permitting lower pipeline velocities without choke-ups. Exact data is not available to evaluate settling rates. For a factor of 100 a suspension of clean building sand in water is suggested. When pumping long distances, a quick settling matrix allows the coarser solids to settle out along the bottom of the pipeline, causing drag, turbulence, and increased friction. With a slow settling matrix as at Noralyn, turbulence acts to keep the solids in suspension at a lower friction head, regardless of the pumping distance. When the pebble, content of the matrix, i.e., the +14 mesh fraction, is in excess of 10 pct of the total solids, trouble may be expected from settling out even in normal pumping distances. To prevent choke-ups and maintain tonnage, an additional pump must be added in the long runs, where one pump would otherwise be satisfactory. A typical pulp handled is: total volume, 7800 gpm; water, 4500; solids pumped per hr, 4200 lb; sp gr pulp, 1.4; percent solids in pulp, 46.; pipe size, 16-in. ID; pulp velocity, 12.85 fps; probable critical velocity, 10 fps, as below this minimum -velocity choke-ups would be numerous. In calculating friction heads the Armco handbook is used where a roughness factor based on 15-year-old pipe is set up. Because the pipe used in pumping matrix is, smooth and- polished because of the scouring, action of the phosphate and its silica content, the head losses in the Armco table for water are practically the same as in pumping the Noralyn matrix through smooth pipe, plus the fact that conditions vary widely over short periods, making accurate determinations difficult to obtain. New pumps and pump, changes are being tested continuously and a wealth of data built up. This has resulted in a substantial improvement and lower relative costs in pumping matrix. The Florida phosphate industry is constantly seeking to offset higher wage and material costs with improved technique. Until a few years ago a 12-in. discharge pump was commonly used, with heads as low as 80 ft. Sizes have gradually increased and heads more than doubled. For example, the following pump was placed under test at the Noralyn mine: make, Georgia Iron Works; size, suction 16 in., discharge 14 in.; impeller, 39-in. diam; motor, 600 hp, slip ring; full load speed, 514 rpm. The results were increased head, higher capacity than the older design, with fewer pumps in the line from mine to washer.

Jan 1, 1952

By D. M. Boulin, E. S. Greiner

Single crystals of vanadium were bent at 293" and 77°K to a maximum fiber strain of about 1 pct. Analysis of slip line and etch markings of- crystals deformed at either temperature show the slip planes are {112}. The slip directions at 298° and 77°K, deterwined from the asterism axes of rotation of crystals compressed in a {110} direclion, were (111). Etch pits, associated with dislocations, were observed in the plastically deformed vanadium. There was no evidence of mechanical twinning. THIS paper presents results on single crystals of vanadium plastically deformed by bending and compressing at 298° and 77°K. Bcc metals have been observed to slip on {110}, (1121, and (123) planes while the slip direction has invariably been found to be ( 111) 1 Lindley and smallman2 have studied the deformation intension of polycrystalline vanadium between 298° and 20°K and reported that at and above 77°K the deformation is predominantly by slip, but at 20°K the deformation is associated with twinning. Later, Edington and smallman3 found that single crystals of vanadium compressed in a (111) direction at 77°K deformed by twinning on { 112}( 111) systems. Clough and pavlovic4 have reported that mechanical twins were observed in vanadium loaded by impact at 77°, 173°, and 195°K, respectively. PREPARATION OF SINGLE CRYSTALS The vanadium single crystals were made from commercial vanadium reported to contain 0.03 pct C, 0.03 pct 0, 0.02 pct N, 0.0X* pct Ca, 0.0X pct Si, proximately 0.8 cm diam, 4 cm long and were grown in (100) and (110) directions. A crystal for studies of etch pits was gound to have a face parallel to a {l00} plane by the Bond crystal-orienting technique.7 In order to avoid straining, the crystals were placed on the steel mount with a low-melting organic compound (glycol phthalate) before grinding. Crystals for the bending experiments were ground to have the four side faces parallel to ( 100) planes. The crystals were electropolished8 in 10 pct H2So4 using a current density of about 3 amp per sq cm to remove the strained surface material. The crystals were bent slowly by three-point loading in a stainless-steel jig designed by Sylwestrowicz.9 The end rollers of the jig were 3.8 cm apart. DEFORMATION AT 298°K The bent vanadium crystal, after removal from the jig, had a radius of curvature of 20.4 cm over the 2-cm length close to the center loading roller, where the maximum bending moment occurred. This radius of curvature corresponded to a maximum fiber strain of 0.9 pet. The side faces contained slip bands as shown in Fig. 1. These bands are straight where the strain is high, top and bottom of micrograph, but are curved on either side of the neutral axis in the center. This curvature of the slip bands in the low-stress area may be due to cross slip. The plastically deformed crystal was subsequently electropolished and then electroetched in 10 pct H2SO4 solution, using a current density of 0.25 amp per sq cm. The structure is shown in Fig. 2. The directions of the traces on two adjoining faces were plotted stereo graphically, Fig. 3.* Since there were two sets of

Jan 1, 1968

By Fred C. Bond

Many grindability tests have been made in the Allis-Chalmers laboratory since the tabulated results were last published.&apos; The lists of standard ball-mill and standard rod-mill grindability tests are brought up to date in this paper, and considerable new material is added. This includes a tabulation of open-circuit grinding tests, with impact and compression-crushing tests. The standard ball-mill grindability tests (Table I) were made in a standard laboratory ball mill, 12 by 12 in. inside, running at 70 rpm, with a charge of 285 iron balls ranging from 1½ to 3/4 in. in diameter and weighing 20,125 grams. Seven hundred cubic centimeters of minus 6-mesh stage-crushed dry feed was used, and the circulating load was maintained constant at 250 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 28 to 200, and the grindability is the number of net grams of screen undersize produced per revolution. The table includes 479 tests. It has been calculated that the ball mill does 52.3 joules of useful work on the ore per revolution2 with 93 joules input to the mill. The standard rod-mill grindability tests (Table 2) were made in a laboratory tilting rod mill, 12 in. in diameter by 24 in. long inside, with a wave-type lining, running at 46 rpm. The grinding charge consisted of six I1/4-in. diameter and two 13/4-in. diameter steel rods, all 21 in. long and weighing 33,380 grams. The mill was rotated level for eight revolutions, then tilted up 5&apos; for one revolution, down 5° for one revolution, and returned to level for eight revolutions. It was discharged by tilting downward at 45° for 30 revolutions. Feed consisted of 1250 cc of minus in. stone and the circulating load was maintained constant at 100 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 3 to 65, and the rod-mill grindability is the number of net grams of screen undersize produced per revolution. The table includes 105 tests. The impact and compression-crushing tests (Table 3) are given in a table that is an extension of a previous publication8 in which the method of testing was described in detail. Two equal hammers weighing about 30 lb each were arranged to strike simultaneous blows on opposite sides of the test specimen. Broken pieces of rock passing a 3-in. square and retained on a 2-in. square were set up with the smallest dimension between the hammers, and this distance was measured. The height of the hammers was increased slightly before each blow until the specimen broke, and the foot-pounds per inch of thickness was calculated. Ten pieces

Jan 1, 1949

By John Mullen

Impingers TWO types of midget impinger apparatus are used at Climax. The Mine Safety Appliances Company&apos;s midget impinger (MSA) and the Climax Molybdenum Company&apos;s midget impinger (CMC) (Fig. I). The Leiman Brother&apos;s No. 25, positive displacement pump is used in the CMC impinger instead of the four-cylinder radial pump that is used in the MSA impinger. With the MSA apparatus it is possible to obtain the proper vacuum by adjusting the needle valve on the vacuum chamber, while with the CMC impinger the proper vacuum is obtained by varying the rate at which the pump is cranked. The Leiman Brother&apos;s No. 25 pump is built to displace 0.1 cu. ft. of air per min. when operated at 145 r.p.m. In the CMC impinger this pump is so geared that the required crank speed is approximately 59 r.p.m. Both types of impinger apparatus are equipped with a water gauge mounted on top of the case and they are calibrated and operated in reference to the gauge. The MSA impinger flasks and nozzles are used with both types. Calibrating Impingers A calibrated precision Sargent wet-test meter (Fig. 2) is used in calibrating the impinger apparatus. The apparatus is set up as for taking a sample in the field, except that the hose from the outlet side of the meter is attached to the impinger nozzle. One revolution of the large hand of the meter indicates 0.1 cu. ft. of air displaced. The impinger pump is cranked for one minute, therefore, at a rate that will cause the large hand of the meter to make one revolution. The gauge reading is noted, then the impinger is cranked at this rate for 10 min. and is checked for error. If the meter indicates that more or less than I cu. ft. of air has been displaced, the impinger must be cranked—ither slower or faster according to the error—for another 10 min. and again checked for error. This operation is repeated until the proper gauge reading for a sampling rate of 0.1 cu. ft. per min. has been determined. This gauge reading is used in all sampling operations until the impinger is again calibrated. An impinger is calibrated at least twice each week. MlCROPROJECTOR A Bausch and Lomb model B micro-projector is used (Fig. 3). It consists of a substantial metal base supporting the light source and a microscope. Illumination is provided by a Bausch and Lomb 4.5-amp. arc lamp with mechanical feed. A glass water cell filled with a heat-absorbing and color-filter solution is placed between the arc lamp and the microscope. Two types of solution have been found satisfactory— a 2 per cent solution of copper sulphate or a 2 per cent solution of picric acid and distilled water. This percentage may be varied to give the desired color to the light reflected upon the screen. Any standard microscope may be used. A Bausch and Lomb microscope with

Jan 1, 1946