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By Miles L. Silberman, Edwin H. McKee, John H. Stewart

Three, major suites of igneous rocks were erupted in the Great Basin in middle and late Cenozoic time. The first resulted in eruption of andesitic rocks in the northern Great Basin between 42 and 34 m.y. ago. Between 34 and 17 m.y. ago this igneous activity migrated to the southwest, ending about 17 m.y. ago with the onset of crustal extension and north-south- trending Basin and Range (rift) faulting. The second volcanic suite erupted between 17 and 8 m.y. ago in the northern Great Basin. The third volcanic suite consists of andesitic and rhyolitic lavas and some basalt erupted between 22 to 7 m.y. ago in the western Great Basin along the northwest transcurrent Walker Lane fault zone. Hydrothermal ore deposits are associated with all suites of igneous rocks.

Jan 1, 1977

INTRODUCTION Conditions for formation of porphyry copper deposits appear to have been propitious in the Appalachian orogen from the end of Precambrian into Middle Ordovician, and again from Middle Devonian into Mississippian. In the Appalachians we can now record the evolution of a porphyry copper province within a developing orogen, one that reached metamorphic and structural maturity by the end of Devonian. Fig. 15, after Irwin and Coleman (1972), locates that portion of the Appalachian orogen included in this study, and Fig. 16 illustrates general geology of the area. Deposits mentioned in Table 3 and in the text are shown in Fig. 16. Data presented summarize personal observations by the author, published descriptions of specific areas noted in the bibliography, and abstracts from Provincial assessment files. Sources of previously published data are acknowledged. Few comprehensive descriptions of individual porphyry copper deposits in this area exist, so much of the data presented is new to the literature. Porphyry mineralization has, for this chapter, been restricted to porphyry copper de- posits, although porphyry molybdenum de- posits (Clark, 1972) are also shown in Fig. 16 (Hollister, et al., 1974). Copper- molybdenum and tungsten-molybdenum skarns are also omitted to simplify discu sion, though each is known in the area studied. Gaspe Copper is included, however; min- eralization in this deposit is present in the intrusion as well as in the skarn adjacent (Bell and Scott, 1954). Porphyry deposits listed in Table 3 are included because of some important feature, although other deposits have been reported (e.g., Kirkham and Soregaroli, 1975 ) . Clearly, no table now compiled could be complete. The pace of exploration insures that new occurrences will probably be found. Summaries of features exhibited by intrusions shown in Table 3 should remain valid, however, since they appear representative and are based on an adequate number of deposits in the area studied. Paleozoic porphyry copper prospects (shown in Fig. 16) are more common, apparently, in the northeastern part of the orogen, rarely having been found south of Catheart or west of the Logan Line. Occurrences of molybdenum and other associated lithophile elements are, however, known to the south and west as well as further north. Killeen and Newman (1965), King (1970),

Jan 1, 1978

By E. L. Kamen, H. D. Kessler, M. Hansen, D. J. McPherson

The highly reactive Ti-Mo and Ti-Cb alloys were prepared and heat treated under protective conditions. Phase diagrams were established based on micrographic and X-ray diffraction analysis and detection of incipient melting. ß titanium forms a continuous series of solid solutions with both molybdenum and columbium, whereas these metals are only slightly soluble in the a titanium modification. PREVIOUS work on the Ti-Mo and Ti-Cb systems was very limited and is discussed in succeeding sections of this paper relating to the respective phase diagrams. The diagrams in the present work were determined principally by micrographic analysis, because other methods such as thermal analysis and dilatometry are not suitable, due to the low diffusion rates of these alloys in the transformation range. X-ray analysis was believed less useful than micrographic analysis because: 1—On quenching certain high titanium alloys from above their transformation range, the body-centered cubic ß phase decomposes into acicular a (hexagonal close-packed) which room temperature X-ray diffraction techniques cannot differentiate from isothermal a. 2— High temperature X-ray techniques might be used, but are difficult to apply and are subject to error because of the high reactivity of titanium with gases such as oxygen and nitrogen. Although metallo-graphic study was the basic method used, X-ray diffraction analysis was applied to a limited extent, to further substantiate the findings. The transformation ranges on the titanium side of the Ti-Cb and Ti-Mo systems were studied both with DuPont Process A (magnesium-reduced) titanium alloys and with iodide titanium alloys. The Process A metal was used chiefly to bracket the transformation range in order to reduce the number of alloys given a particular treatment for the more accurate final study with iodide titanium alloys. Experimental Procedure The analyses of the metals used in the preparation of alloys for this study are presented in Table I. Melting Practice: Early in the work, a consumable electrode, arc melting furnace was used in the preparation of the Process A Ti-base Mo and Cb alloys. This type of furnace has been described by other investigators.', ' One of the major difficulties of the consumable electrode melting process was the prep- aration of homogeneous electrodes (with the mixture of coarse titanium sponge and fine powder alloying addition) which would give homogeneous ingots. Composition gradients encountered may have been due to a tendency for powdered alloying metals to drop away from the electrode in handling. This difficulty in the preparation of homogeneous alloys was overcome by the use of a nonconsumable electrode furnace for later work. This furnace is similar to the one used by Mc-Pherson and Fontana a Ohio State University. A diagram of the nonconsumable electrode arc furnace is given elsewhere.' The source of power was a 400 amp dc welding generator. Electrode tips of tungsten (% in. in diam) were used for all columbium melts and for a limited number of molybdenum melts. Molybdenum tips were used for the majority of molybdenum alloy melts. Columbium tipped electrodes were tried, but disintegrated rapidly at currents above 300 amp. In the Process A titanium melting practice the metals were charged into the water-cooled spun copper crucible in the following forms: l—titanium: sponge (—4 + 16 mesh); 2—molybdenum: sections of 0.003 in. sheet (approximately ½ in. sq); and 3— columbium: powder compacts (? in. lumps broken from large section). After alternately evacuating and purging the furnace several times with argon or helium, the arc was struck against the metal charge,

Jan 1, 1952

By E. H. Mohr

At the McGill concentrator of the Nevada Consolidated Copper Co., all flotation operations have been carried out in Forrester cells since November, 1926. In respect to cost of operation, the new cell has proved to be superior to the Callow, which it supplanted, and its incorporation in the flow sheet has resulted in a marked improvement in the metallurgy. Before citing the advantages of the new machine over the old, it should be explained that the Forrester has not only replaced the Callow, but it has made possible the elimination of a number of tables as well. When Callow cells were in use, table concentration preceded flotation. The ball mill pulp was elevated, by means of air lifts, to roughing Wilfley tables situated two floors directly above the flotation floor. On an intermediate floor were additional tables which cleaned the concentrate from those above. The finished table concentrate constituted approximately 60 per cent. of the total mill production, and contained about 40 per cent, of the recovered copper. The table tailings were returned to the classifiers, the overflow from which went to the flotation cells and the oversize to the ball mills. As soon as the Forrester cells were installed, this tabling operation was discontinued. The ball mill pulp now discharges into the classifiers direct, without the intermediate tabling. It was impracticable to ernploy the present flow sheet when Callow cells were in use, for the reason that unless the great,er part of the heavy mineral was first removed, it accumulated on the mats and adversely affected the flotation operation. By tabling ahead of flotation, it was practicable to maintain the concentrator at 10 to 20 per cent. greater capacity (depending up on the pyrite contnt of the ore) than was possible by flotation alone. This was due to the necessity for finer grinding in the latter case. Even though the table conecntrates contained only 6 to 8 per cent'. copper content, the low smelting costs made it advantageous to operate at high mill capacity, at the cost of low grade of product. Standard Section equipment With the old flow sheet, a standard section consisted of four ball mills and classifiers, one air. lift, 20 roughing tables, 10 cleaner tables, 12 primary

Jan 1, 1928

By KING HAMILTON GRAYSON

IRON and steel were the only basic industries in the Soviet Republic in 1928 that lagged behind the pre-war production on a comparative basis. This was due to the almost complete obliteration of all industrial units manufacturing iron and steel products during the Revolution, which resulted in the production of 1921 being only 4 per cent of that of 1913. By the end of 1927 the production of pig iron had reached a point equal to 70.5, per cent of 1913; steel was 84 per cent ; and rolled iron was 78 per cent. According to the. planned program of the Supreme Economic Council, the pre-war level of the iron and steel industry was to have been reached and exceeded during 1929-30. The production of the entire period from 1917 to 1928 in comparison with 1913, is shown in the following table

Jan 1, 1929

By Samuel S. M. Chan

The magnetic susceptibilities of some ores and the major rock formations of the Precambrian Belt Supergroup in the Coeur d'Alene mining district were determined both in the laboratory by the use of the Gouy method and in the field with a Bison susceptibility unit. Trace element analysis of the same group of rock samples was also made using the atomic adsorption technique. Results of this preliminary study indicate that some correlation exists between the magnetic susceptibilities and the various rock formations. This may be useful for regional mapping and mineral exploration.

Jan 1, 1974

By Charles L. Mantell, James E. Hoffmann

Mechanical inclusion, electrophoretic deposition, and adsorption were studied as mechanisms for code-position of aluminas present in copper-plating electrolytes as an insoluble disperse phase. Mechanical inclusion was not a significant factor. That codeposi-tzon of aluminas by an electrophoretic mechanism was unlikely was substantiated by measurements of the potential of the aluminas. The alumina content of the deposits was studied as a function of the pH of the bath. These tests in conjunction with sedimentation studies demonstrated the absence of an isoelectric point for the alutninas over the pH range examined. Thiourea in the electrolyte (a substance known to be adsorbed on a copper cathode during electrodeposition) affected the amount of alumina in the electrodeposit. However, no adsorption of thiourea on aluminas in aqueous dispersions was detected. If it were possible to produce a dispersion-hardened alloy of copper and alumina by electrodeposition, an alloy possessing both strength and high conductivity at elevated temperatures might be anticipated. Investigation of the mechanism of codeposition of aluminas with copper was undertaken with the hope that knowledge of the mechanism would aid in the development of such an alloy. The word "codeposit" here does not necessarily imply an electrolytic phenomenon but rather that the materials codepositing, the various aluminas, are transported to and embedded in the electrodeposited copper by some means. Mechanical inclusion in electrodeposition implies a mechanism of codeposition which is wholly mechanical in nature; the only forces acting on a particle are gravity and contact forces. Such a particle is presumed to be electrically inert and incapable of any electrical interaction with electrodes in an electrolytic plating bath. Processes for matrices containing a codeposited phase by electrodeposition from a bath containing a disperse insoluble phase frequently state that code-position is caused by mechanical inclusion.10,2,12 If settling, i.e., gravity, be the controlling mechanism for codeposition of aluminas, then assumptions may be made that 1) the content of alumina in the electrodeposit should be enhanced by increasing the particle size, 2) the geometry of the system, that is, the disposition of the cathode surfaces relative to the di- rection of the falling particles, should affect the alumina content of the electrodeposit, 3) in geometrically identical systems the chemical composition of the electrolyte employed should exercise no effect on the alumina content of the deposit, that is, the alumina content should be the same in all cathode deposits irrespective of bath composition. A bent cathode19 evaluates the clarity of filter effluent in electroplating baths by comparing the roughness of the deposit on the vertical surface with that on the horizontal surface. Two difficulties are inherent in this technique: 1) the current density on the horizontal portion of the cathode would be substantially greater than that on the vertical surface; 2) should the deposit obtained be rough, projections on the vertical face could act as horizontal planes and vitiate the relationship between the vertical and horizontal surfaces. Bath composition should have no substantial effect on the alumina content of the deposit. Two different electrolytic baths were employed. They possessed variant specific conductances and substantially different pH ranges. The experimental tanks were rectangular Pyrex battery jars 6 in. wide by 3 1/4 in. long by 9 3/4 in. deep. The cathodes were stainless steel 316 sheet of 0.030 in. thickness, cut to 7.5 by 1.75 in. and bent at right angles to form an L-shaped cathode whose horizontal surfaces measured 1.75 by 3.0 in. All edges and vertical surfaces were masked with Scotch Elec-troplaters Tape No. 470. The anodes were electrolytic cathode copper 9 in. high by 2.25 in. wide by 0.5 in. thick. To eliminate inordinately high current densities on the projecting edge of the cathode, the anode was masked 1 in. above and below the projected line of intersection of the cathode with the anode. The exposed area of the anode was equal to that of the cathode, providing both with equal average current densities. The agitator in the cell was of Pyrex glass and positioned so its center line was equidistant from cathode and anode, and a plane passed horizontally through the center of the blade would be located equidistant from the bottom of the cathode and the bottom of the deposition tank. The assembled apparatus is depicted in Fig. 1. Hatched areas on anode and cathode represent the area of the electrodes wrapped with electroplaters tape. MATERIALS The chemicals were copper sulfate—CuSO4 • 5H2O— technical powder (Fisher Scientific Co.). Spectro-graphic analysis showed substantial freedom from antimony, arsenic, and iron. Traces of nickel were present.

Jan 1, 1967

By M. E. Grimes

Rio Algom Mines Ltd. owns eight uranium mines and mills in the Elliot Lake district of Ontario, seven of which were producing before 1960. At present only one mine is operating and meeting current contract requirements. Under present economic conditions, it is not feasible to mine ore of less than 2.0 lb per ton U3O8, although there are very large reserves below this grade, Rio Algom is interested in preconcentration for three possible applications: to extend ore reserves by permitting use of lower grade ore, to reduce costs on processing of present grade ore, and to reduce capital investment in expanding production to meet future production goals. Rio Algom has been investigating various applicable preconcentration processes since the early 1950s to date, on laboratory and pilot-plant scale. Methods investigated include gravity concentration by sink-float, high intensity wet magnetic separation, flotation of uranium, flotation of acid consumers, and a number of electronic sorting methods including radiometric, photometric, and conductimetric. No method tested has yet been satisfactory under Rio Algom's criteria of uranium recovery and cost, principally because uranium loss has been too high for a useful weight rejection. Metallurgical results of the preconcentration test work and the economic assessment therefrom are reviewed.

Jan 1, 1974

By J. A. Barr, R. B. Burt, 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 +I50 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, 1953

By R. B. Burt, J. 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 +I50 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, 1953

By K. V. S. Sastry

INTRODUCTION Flotation is by far the most important unit operation of mineral processing. Most of the common base metals are produced from flotation concentrates of metal sulfides. Flotation is proving to be a valuable process for producing non-metallic minerals such as ceramic oxides, silicates, clays and phosphate fertilizers. All these fltation separations are carried out on materials which are predominantly in the 10 to 100 micronssize range. The present practice of flotation circuits for the production of large tonnages of metallic and non-metallic mineral concentrates is fairly well understood, developed, and standardized. Two kinds of problems are encountered (1, 2) with fine particles smaller than 10 microns. First is the interference of the fine particles with the flotation of the coarser particles and the second is the poor flotation behavior of fine particles. In general, these problems are handled either by eliminating, minimizing, or ignoring the fines in the flotation circuit feed. In recent years, however, a considerable number of experimental and theoretical studies have been carried out to understand the flotation behavior of fine particulates. This literature is summarized in four review papers (2-5). Application of the conventional flotation techniques for the separation of mineral fines is known to be uneconomical, and to present several scientific and technological problems (1-5). The objective of this presentation is to examine the flotation behavior of fine particles in terms of their physical-chemical properties and to review the potential techniques of fine particle flotation. Specific reference to the problem areas and research needs will also be included in the discussion.

Jan 1, 1979

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 James K. Stanley

When a plastically deformed mctal is heated to a certain temperature, it undergoes a complete change in microstructure, the consequence of which is a marked alteration of mechanical properties such as hardness, strength and ductility. This change in microstructure occurs by re-crystallization and is a process involving the formation and growth of strain-free nuclei in the deformed matrix. These nuclei continue to form with time and grow until a new grain structure replaces the cold-worked structure. The formation of the nuclei in the deformed sheet of I per cent silicon ferrite is illustrated in Fig. I. If a recrystallization is conducted iso-thermally, a nleasurement of the fraction of the matrix recrystallized plotted against time leads to an isothermal recrystallization curve, first established in 1942.1 Such recrystallization curves are shown in Fig. 4. These curves show that the recrystallization of a cold-worked metal is best considered as a rate process rather than a time independent process such as is suggcsted by the improperly named and misleading recrystallization diagrams involving deformation, temperature, and grain size.2-5 Any study of the rate of recrystallization resolves itself into a study of the component rates of nucleation and growth. The previous paper1 provided an account of the determination of the rates of nucleation; N, the number of nuclei formed per unit time per unit of unre-crystallized area, and growth, G, the radial growth in unit time, in the recrystallization of silicon ferrite (I per cent Si). A method was developed to furnish experimental data for evaluating the effect of the recrystallization variables on N and G. The mathematical analysis for the derivation of the isothermal recrystallization curve was developed for a two-dimensional case (in thin sheets where the diameter of the recrystallized grain is greater than the sheet thickness). The present paper is a continuation of the study of recrystallization of I per cent silicon ferrite begun in 1942.1 In this work the effect of the factors deformation, temperature, grain size, composition, and recovery is studied with respect to the rates of liucleation and grouth. At the same time it is shown how the cooperation of N and G affects the isothermal recrystallization curve. The study of these variables on N and G has a definite bearing on the mechanism of recrystallization and should lay a foundation for a rational theory on the subject. Procedure The procedure used in these experiments was essentially the same as that described previously but in certain details some changes were made in the interests of speed. With few exceptions, the procedure is the same for the study of all the variables. Where the procedure is different from standard manipulations attention is called to this fact.

Jan 1, 1945

By R. I. Jaffee, J. W. Roberts, D. N. Williams

DISPERSION hardening as an alloying process has aroused increasing interest in the past few years. This alloying procedure, in which an insoluble phase is dispersed randomly through a metal or alloy, is believed to be especially suited for the attainment of creep strength at elevated temperatures since the dispersed phase is metallurgically stable. Another advantage is that the physical properties of the matrix phase, such as electrical conductivity, remain relatively unaffected by alloying. These two characteristics resulted in interest in the applicability of dispersion hardening to copper and copper alloys. A study of copper alloys prepared by powder metallurgy methods has shown some strengthening from oxide dispersions.' However, powder metallurgy processing would not be economical for large scale production of dispersion hardened copper-base alloys. Therefore, the research described in this article was initiated to examine methods of forming dispersion alloys which would not require powder metallurgy processing. If powder metallurgy processing is to be avoided, a method must be found by which the dispersion material can be introduced into and dispersed through the melt before casting. When the dispersed phase is an oxide or similar nonmetallic, it is extremely difficult to form a satisfactory liquid: solid dispersion since successful mixing requires that the melt wet the dispersed phase. It is almost impossible to force a nonwetted particle beneath the melt surface. This difficulty can be avoided, however, if the oxide particles are produced in molten copper by reaction within the melt. The problem would then reduce to maintaining randomness until solidification. Four procedures were considered for producing an oxide particle within molten copper by reaction. In all cases, the intent was to react oxygen (0) with a reactive metal (X) to form an oxide (XO) within the copper melt. These reactions were as follows: 1) (Cu-X), + (O)g (Cu)1 + (XO)S [1] 2) (Cu-X)1 + (Cu2O)s (Cu)x + (XO)S [2] 3) (Cu-OK)1 + (Cu-X)1 (Cu)1+(XO)9 [3] 4) (Cu-OK)1 +(X)s (Cu)1 +(XO)s [4] The reactions were carried out in an 8-lb melt of 99.98 pct Cu maintained in the vicinity of 2200oF by induction heating. Where possible the amount of the reactants was controlled to give 1 to 5 vol pct oxide. Mechanical stirring during the reaction period was used to minimize agglomeration of the reaction product. The presence of a reaction product was determined metallographically by study of an ingot cast from the melt at the end of the reaction period. Reaction [I] was studied by dissolving aluminum in the melt followed by forcing oxygen through the melt. The melts became quite sluggish during the reaction period suggesting that A12O3 was formed. However, the reaction product rose out of the melt rapidly, possibly due to the flushing action of the oxygen stream, and little dispersed phase was formed. Reaction [2] was examined by dissolving aluminum, silicon, or nickel i; the melt and then adding Cu20 (melting point 2250F). This procedure was unsuccessful. Although Cu2O was easily stirred into cop-

Jan 1, 1961

By Yuri D. Tretyakov, Robert A. Rapp

The stoichiometry ranges ofNiOl+y and LiFe,O,-d were established by high-temperatwe electrochemical meas7rements in a stabilized-zirconia electrolyte cell. The results were consistent with doubly ionized cation vacancies in NiO,+y and interstitial lithium or iron ions in . The defect structure of the ternary ferrite was derived from the consideration of equilibration with respect to oxygen between the solid and the gas phase. The absolute magnitudes of defect concentrations were calculated. Pavtial molar enthalpies of oxygen in the compounds were calculated and interpreted in terms of the enthalpy of defect formation in these crystals. NICKEL oxide (NiO,,?) is a metal-deficient, p-type, extrinsic semiconductor whose properties are consistent with a structural model based on the presence of cation vacancies as the predominant ionic defect at sufficiently high temperatures and oxygen activities. A survey of previously reported conductivity studies and the presentation of some more recent conductivity measurements will be given in a later paper.' The absolute magnitude of the equilibrium vacancy concentration in NiO has been reported from combined conductivity and thermogravimetric data of ' However, disagreement exists concerning the state of ionization of the nickel vacancies in NiO. Some authors3-' have proposed that the predominant defects in NiO are singly ionized nickel vacancies and positive holes (h') formed by the reaction where, accorhng to the notation of Kroger and Vink,6'7 Oq represents an oxygen ion on its normal lattice site. Other authors"27E have proposed that doubly ionized nickel vacancies (V{i ) and positive holes are predominant and are formed by the reaction One purpose of the present investigation was to establish the nature of the predominant defect in NiO,+? as well as its equilibrium concentration and thermo-dynamic properties at elevated temperatures and known oxygen activities. To accomplish this purpose, the coulometric titration of oxygen into and out of NiO was accomplished using a galvanic cell involving the calcia-stabilized zirconia electrolyte. The ternary oxide, lithium ferrite (LiFe50,-6) is ferromagnetic and has the inverse spinel structure LiFe,O,-d Thus, in the ideal stoichiometric XY204 lattice, Fe3 ions occupy one-eighth of the tetrahedral (A) sites, and i' and Fe3' ions share at random one-half of the octahedral (B) sites within the fcc sublattice of oxygen ions.g The structure and thermodynamics of the spinel structures have been comprehensively decribed.'-' A recent compilation of literature for lithium ferrite is also available. 15 Oxygen-excess LiFe50sis not expected to exist because both i' and 17e3& ions are in their highest normal valency states (positive hole formation is not favorable). As will be discussed, equilibration of the ferrite crystal with oxygen of a surrounding gaseous phase will result in the introduction of equilibrium concentrations of ionic and electronic defects. In the Results and Discussion section a defect model for LiFe50,-d is proposed. This defect model is tested by high-temperature coulometric titration experiments. EXPERIMENT The oxygen activities in nonstoichiometric NiOl+, (and also LiFe,O,-d) were measured by means of coulometric titration with the high-temperature galvanic cell The critical characteristic of a suitable experimental cell is the complete isolation of the phase to be investigated in a minimum sized chamber which is free from extraneous sources and sinks for oxygen (leakage). Then oxygen is only admitted to the chamber or removed from it in known amounts by coulometric titration, which involves the passage of oxygen ions through the solid electrolyte with electrochemical oxidation and reduction reactions at the platinum contacts to the electrolyte. The experimental cell is shown in Fig. 1. The cap of the cell was the tip from a closed-end alumina tube, which was found to be leak-free from a helium leak-detector test. This alumina cap was about 1.2 cm OD and about 1.5 cm high. A Zircoa calcia-stabilized zirconia tablet (crucible lid) of 1.5 cm diam and 0.3 cm thickness served as the solid electrolyte. A Pyrex ring of about 0.1 cm thickness was placed between the electrolyte and the cap. The electrolyte tablet was painted with platinum paste on the entire outer face and on that part of the inner face which would be within the enclosure; these electrodes were further prepared by heating in air at 1000°C for 2 hr. Nickel oxide powder, listed as 99.999 pct pure, was purchased from Leico Industries, Inc. Pills of the NiO were cold-pressed and sintered at 1050°C for 3 hr in a platinum crucible. This poorly sintered NiO was crushed, and chunks were wrapped in 52-mesh Pt gauze (to catalyze the solid-gas exchange) for place-

Jan 1, 1970

By James W. Dally, William L. Fourney, Anders Ladegaard Peterson

For the past 3 years, the authors have been conducting research sponsored by the National Science Foundation (RANN) to improve the process of excavation by drilling and blasting. The approach followed has been experimental where the development of stress waves and fractures initiated at the bore hole have been investigated in order to obtain a complete understanding of the dynamic Fracture process. The second step in the approach has been to introduce modifications in the drill and blast procedure which will permit closer control of the fracture process.

Jan 2, 1978

By R. C. Corson

Modern microprocessor systems make it possible to use several variables to anticipate changes in the process and to react with responses before upsets can be magnified by the process or by control action. Through tighter control, "insurance" can be eliminated from process set points without affecting product quality. In actual operation with a traveling grate furnace this control philosophy has produced savings of up to 20% in fuel consumption and 8% in fan power usage while producing a more consistent quality product.

Jan 1, 1984

By N. J. Themelis, A. W. Last

Autogenous grinding, broadly defined as the self-comminution of ore in a cylindrical tumbling mill without using auxiliary grinding media, was first applied in the early 1930s in the so-called Hadsel mills (Hardinge, 1955) as seen in [Fig. 1], mainly used in gold mines on the US West Coast. A number of these mills were built in sizes of 6.1 to 7.3 m (20 to 24 ft) in diameter and 1.5 to 1.8 m (5 to 6 ft) in length but were later abandoned because of operating and maintenance problems. Further development of autogenous grinding led to the Hardinge wet grinding mill, the Aerofall dry grinding mill, and others. A tabulation of all known autogenous and semiautogenous installations around the world was sent recently to the authors by G. J. Kushner of Koppers Metal Processing Division. In the size range of 7.3 to 9.8 m (24 to 32 ft) in diameter, it lists a total of thirty mills used in the processing of copper ores. An additional 120 units are listed in the same size range for the processing of iron ores and other minerals. The largest mills in existence are 11.1-m-diam X 4.6-m-long (36-ft-diam X 15-ft-long) mills in Minnesota, powered by 8950 kW (12,000 hp) motors and used in taconite beneficiation. It is indicative of the phenomenal growth of this technology that none of the above 150 mills was in operation prior to 1964. Grinding rates in autogenous mills are very sensitive to variations in ore breakage characteristics and even with fairly constant ore type can exhibit large fluctuations in throughput. Autogenous grinding also depends on availability of a fraction of the ore to be ground in the form of "competent" rock which can withstand to a certain degree the tumbling and impact action within the mill. If the ore is uniformly "soft", it will soon be crushed into small size particles and further grinding, which depends on the existence of large size rocks in the mill, becomes impossible. This apparently is the case with most copper ores, as discussed by Bachman, Last and Nabbs (1970) who described pilot testing of a large number of ore samples from Kennecott 's Chino Mines. Under circumstances where autogenous grinding is impractical, it is possible to realize most of its benefits by utilizing large grinding balls in place of the "competent" rock and practice what has become known in the mining industry as semiautogenous grinding. In semiautogenous grinding (SAG), a grate discharge tumbling mill is fed the product of a primary crusher or run-of-mine rock. Forged steel or alloy grinding balls, in proportions ranging from 3.10% of the total volume in the mill, are used to complement the self-grinding ability of the ore. Semiautogenous circuits can be either designed for single-stage closed-circuit grinding which produces a final product for recovery processes, or for two-stage operation in which the SAG mill product is fed to a second stage of grinding in a conventional ball mill.

Jan 1, 1981

By Martin A. Mayers

It has long been recognized that it would be highly desirable to be able to predict the temperatures at various points in a burning fuel bed and their variations with changes of the properties of the fuel and the rate of air supply. Experimenta1 investigations of fuel-bed temperature1-4 have not been lacking, but the absence of any theoretical background has made it impossible to use the results of these tests as a basis for prediction, except in a very general may. A mathematical analysis", based on a somewhat simplified theory of the distribution of lieat in a fuel bed, lias recently been completed by the author. Even though numerical values of some of the coefficients are not known, the calculations result in some conclusions of general significance. Thus, for example, it appears that when the temperatures throughout the fuel bed can be calculated, the rate at which ignition of the fuel by underfeed action can be maintained can also be calcu1ated. The uncertainty concerning the values of the coefficients referred to above is serious limitation on the use of that method. The following paper presents a discussion of these coefficients and includes estimates of their values based on the data now available. A brief description of the theory on which the analysis is based, and some results of the calculations showing their degree of agreement with experimental results, are presented by way of introduction to the discussion of the coefficients. For a more. extended discussion of the theory, see the A.S.M.E. paper5. Theory and Comparison with Experiment The theoretical analysis is similar to that used in the calculation of the stationary temperatures in a conducting solid in the classical theory of heat concluction. The fuel bed is pictured as a uniform, continuous, porous solid through the interstices of which air can be blown and ill which combustion can take place. The solid lias a finite thermal conductivity, so that heat may be transported in any direction, against the flow of air as well as in the same direction. While it is true that nothing

Jan 1, 1938

By Martin A. Mayers

It has long been recognized that it would be highly desirable to be able to predict the temperatures at various points in a burning fuel bed and their variations with changes of the properties of the fuel and the rate of air supply. Experimenta1 investigations of fuel-bed temperature1-4 have not been lacking, but the absence of any theoretical background has made it impossible to use the results of these tests as a basis for prediction, except in a very general may. A mathematical analysis", based on a somewhat simplified theory of the distribution of lieat in a fuel bed, lias recently been completed by the author. Even though numerical values of some of the coefficients are not known, the calculations result in some conclusions of general significance. Thus, for example, it appears that when the temperatures throughout the fuel bed can be calculated, the rate at which ignition of the fuel by underfeed action can be maintained can also be calcu1ated. The uncertainty concerning the values of the coefficients referred to above is serious limitation on the use of that method. The following paper presents a discussion of these coefficients and includes estimates of their values based on the data now available. A brief description of the theory on which the analysis is based, and some results of the calculations showing their degree of agreement with experimental results, are presented by way of introduction to the discussion of the coefficients. For a more. extended discussion of the theory, see the A.S.M.E. paper5. Theory and Comparison with Experiment The theoretical analysis is similar to that used in the calculation of the stationary temperatures in a conducting solid in the classical theory of heat concluction. The fuel bed is pictured as a uniform, continuous, porous solid through the interstices of which air can be blown and ill which combustion can take place. The solid lias a finite thermal conductivity, so that heat may be transported in any direction, against the flow of air as well as in the same direction. While it is true that nothing

Jan 1, 1938