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By N. A. Gokcen, J. Chipman

Aluminum and oxygen dissolved in liquid iron were brought into equilibrium with pure alumina crucibles and atmospheres of known H2O and H2 contents to study the reactions: 1—Al2O3(s) = 2 Al + 3 0; 2—Al2o3(s) + 3H2(g) = 2Al+ 3H2o(g); and 3—H2(9) +O = H2O(g). Aluminum strongly reduces the activity coefficient of oxygen and similarly oxygen reduces that of aluminum. Values of the product [% All" • [% O]3 are much smaller than those found in previous experimental studies and are of the order of magnitude of the calculated values. ALUMINUM is the strongest deoxidizer commonly A used in steelmaking, but the extent to which it removes dissolved oxygen has been debatable. The relationship between aluminum and oxygen has not been determined reliably not only on account of the usual experimental difficulties at high temperatures but also because of uncertainties in the analyses of very small concentrations of oxygen and aluminum. The earliest experimental attempt of Herty and coworkers' was followed by a more systematic study of Wentrup and Hieber.' These authors added aluminum to liquid iron of high oxygen content in an induction furnace and considered that 10 min was sufficient to remove the deoxidation products from the melt. Parts of the melts thus obtained were poured into a copper mold and analyzed for total aluminum and oxygen (soluble plus insoluble forms), assuming that the insoluble parts were in solution at the temperatures from which samples were taken. It is conceivable that the furnace atmosphere in their experiments, consisting of mainly air at 20 mm Hg pressure, was a serious source of continuous oxidation and therefore that their oxygen concentrations were correspondingly high. Scattering of their data was explained to be well within the maximum inaccuracy of 10°C in the temperature measurements and errors of ±0.002 pct each in the oxygen and total aluminum analyses. Maximum and minimum deoxidation values, i.e., values of the product [% All' . [% O] differed by factors of 10 to 15; mean values of 9x10-11 and 7.5x10-9 ere reported at 1600" and 1700°C, respectively. Hilty and Craftsv determined the solubility of oxygen in liquid iron containing aluminum, using a rotating induction furnace. Pure alumina crucibles used in their experiments contained the liquid iron which in turn acted as a container for slags of varying compositions consisting mainly of Al2O3, Fe2O3, and FeO. The furnace was continuously flushed with argon, and additions of aluminum and Fe2O3 were made in the course of each experimental heat. The inner surfaces of their alumina crucibles were covered with a substance other than pure Al2O3, containing both iron oxide and alumina. Although frequent slag additions can change the composition of slag in the liquid iron cup formed by rotation, the inner surface of the crucible must depend upon the transfer of oxygen or aluminum through the liquid iron for any adjustment in composition. It is not clear that their metal was in equilibrium with the crucible wall, but it is clear that it was not in equilibrium with Al2O3. Their deoxidation product, [% A].]" • [% O]3, varied by a factor of more than 50; the average values of 2.8x10- and 1.0x10-7 were selected for temperatures of 1600" and 1700°C, respectively. Aside from the experimental determinations, attempts have been made to calculate the deoxidation constant for aluminum indirectly from thermody-namic data. Schenck4 combined the thermodynamic data for Al2O3 and dissolved oxygen in liquid iron by assuming an ideal solution. His calculated values are 2.0x10-15 and 3.2x10-13 at 1600" and 1700°C, respectively. Later, Chipman5 attempted to correct for the deviation from ideality and derived an expression which led to deoxidation values of 2.0x10-14 and 1.1x10-12 at 1600" and 1700°C, respectively. The errors in these treatments originate mainly from inaccuracies of thermal data and uncertainties regarding the activity coefficients of dissolved oxygen and aluminum. The purpose of this investigation was to study the equilibria represented in the following reactions in the presence of pure alumina: Al2O3(s) = 2Al + 3O K = aAl2.ao3 [1] Al2O3(s) + 3H2(g) = 2Al + 3H2O(g) H2O K2 = aAl2(H2O/H2 ) [2] H2(g) +O = H2O(g) K3 = 1/ao (H2) [3] The experimental method consisted of melting pure electrolytic iron, usually with an initial charge of aluminum, in pure dense alumina crucibles under a controlled atmosphere of H,O and H2 and holding

Jan 1, 1954

By Willard L. Hunter

Four solvent extraction systems were studied to determine their efficiency jor extraction and separation of tantalum and columbium. Aqueous feed solutions of varying HF-HCl concentrations and metal content were contacted with equal volumes of cyclohexanone, 3-methyl-2-butanone, and 2-pentanone and solutions of varying HF-H2S04 concentrations were contacted with equal volumes of 2-pentanone. One multistage continuous test was made in a polyethylene pulse column using cy clohexanone as the organic phase. In each system studied, columbium and tantalum purities in excess of 95 pct with respect to each other were obtained in single-stage tests at low acidities in the feed solution. Separation factors ranging from 1700 to 2400 were obtained when rising HF-HCl mixtures in the aqueous phase. Best results were obtained when a solution of HF-H2S04 was used as the aqueous phase and 2-pentanone as the organic phase. A separation factor in excess of 6000 was obtained in one stage with aqueous solution concentrations of 2 _N HF and 2N H2S0,. When acid concentrations were increaszd to 52 HF and 10 _N H2S0,, 99.9 pct of the tantalum and 98.2 pct of the columbium initially present in the feed solution were transferred to the organic phase. The separation of columnbium and tantalum obtainable by means of the solvent extraction systems presented in this paper was found to corn -pare favorably with other systems, including the HF-H2SO4-methyl isobutyl ketone system currently used by most producers for the extraction and separation of these metals. TANTALUM and columbium are always found together in minerals of commercial significance, although the proportion of the two metals in ores varies within broad limits. Columbium is estimated to be 13 times more abundant than tantalum. Five methods generally employed for the separat:ion of these metals are: 1) fractional crystallization (the Marignac process),2 2) solvent separation, 3) fractional distillation of their chlorides, 4) ion exchange, and 5) selective reduction. Of these methods, the one currently used by industry to the greatest extent is that of solvent separation. One of the early technical developments in solvent separation of tantalum from columbium was reported by the Bureau of Mines: the HF-HC1-methyl isobutyl ketone system; data were presented for both laboratory and pilot-plant experimentation.3 Of twenty-eight organic solvents tested for their ability to extract tantalum from an HF-HC1 solution of columbium and tantalum, 3-pentanone (diethyl ke-tone), cyclohexanone, 2-pentanone, and 3-methyl-2-butanone were chosen for further study. Data on the HF-HC1-diethyl ketone system has been published4 and data describing the use of cy clohexanone, 2-pentanone, and 3-methyl-2-butanone as the organic phase are included in this report. RAW MATERIAL The source of tantalum and columbium oxides for this study was ('Geomines" tin slag from the Manono Smelter, Cie Geomines, Gelges, S.A., Congo. In order to extract the valuable Ta-Cb content, the slags were carbided, chlorinated, and the sublimate from chlo-rination was hydrolized and washed free of chloride with water. The washed material was air-dried and stored in a stoppered container. Throughout the paper, "feed material" refers to this mixture of hydrated oxides which was employed because of its high solubility in aqueous solutions. Typical analysis of the hydrated oxides is shown in Table I. I) HF-HC1-CYCLOHEXANONE SYSTEM Batch Separation. Effect of Acid Concentration. To determine the effect of varying the acid concentration upon the transfer of tantalum and columbium, a series of tests was made in which approximately 2.5 g of feed material was added to 25 ml solutions of 2, 4, 6, 8, and 10 N HF and 0 through 5 N HC1. Tantalum pentoxide concentration of the solu%ons was approximately 21 g per liter and columbium pentoxide was 14 g per liter. These starting solutions were shaken with equal volumes of cyclohexanone in 100 ml polyethylene bottles for 30 min. The phases were carefully separated in 125 ml glass separatory funnels. The time of contact of the solutions with the separatory funnels was kept at a minimum to reduce silica contamination. The measured phases were separated into 400 ml polyethylene beakers and the metal contents of each were precipitated by addition of an excess of ammonium hydroxide. Precipitate from each phase was filtered on ashless filter paper, ignited at 800" to 1000°C for 45 min, weighed, and analyzed by X-ray fluorescence.5 Data tabulated in Table I1 and illustrated in Fig. 1, show that maximum separation of tantalum from columbium for each HF concentration was obtained with no HCl present. The purest tantalum product was obtained with some HCl present. The highest separation factor was obtained at 2 N HF and

Jan 1, 1970

By Robert B. Coleman

The recent surge in demand for uranium in the US, accompanied by a significant increase in price, has stimulated the search for and development of lower grade ore bodies. At the mill, this translates into much larger throughputs and-compounded by many recently enacted environmental considerations-has made the cost of milling a critical factor in today's uranium production economics. The three conventional milling methods employed in the extraction of uranium ores are reviewed in the following sections.

Jan 10, 1978

By J. C. Bierlein, R. A. Dodd

Endurance strengths at 10' cycles, fatigue-hardening rates, and endurance strength/0.2 pct proof stress ratios have been determined jbr a range of Cu(Ni), Cu(Si), and S.R.0. Cu(Au) solid solutions. Some douht is cast on simple cross slip models of fatigue hardening in view of opposite composition dependencies of hardening displayed by Cu(Si) and Cu(Au). The temperature dependencies of hardening also are the opposite of predictions made on the basis of a cross slip model. A correlation apparently exists between the fatigue strength/0.2 Pc~ proof s1.re.s~ ratios and rates of fatigue hardening. An increase in PS/PS due to alloying or temperature change is accompanied by a decrease in hardening rate, and vice versa. In recent years there has been much interest in the determination of the dislocation structure of fatigued metals and alloys. The accumulated evidence, e.g., Refs. 1 to 5, suggests that the structure may be dependent on the stacking fault energy, 7, of the fatigued metal in the same way that fatigue hardening rates have also been shown to apparently depend on u .6' 7 The above research followed the various earlier theories relating crack nucleation to ease of cross slip, e.g., Refs. 8 to 131 so "ggesting that y may indeed be the parameter of principal significance in all facets of the fatigue process. However, comparatively little attention has been paid to correlating engineering fatigue data with observations of the above type. Certainly, it would be useful if potential engineering fatigue performance could be assessed from a knowledge of easily determined alloy properties, and the present research originated from this standpoint. TO investigate the widest Possible range of 7 would require the use of two solvent bases, e.g., aluminum and copper, but a reasonable coverage can be provided by copper-based solutions alone. Therefore, it was decided to work with Cu(Ni) and Cu(Si) solid solutions, the approximate stackillg fault energies of which are given in Table I. The y range extends from low, Cu(7.5 at- Pet Si), to moderately high in Cu(2-5 at. ~ct ~i) and two of the Cu(Ni) alloys. The values listed in Table I should be regarded as qualitative only, being derived as follows. Dillamore and smallman14 quote a value of 85 i 30 erg.cm-2 for the stacking fault energy of pure copper, based on an earlier value due to Howie and swannl5 corrected by Brown's formula.'' Since the ratios (alloy)h(pure copper) have been reported17,18 for Cu(Ni) alloys containing up to 30 pct Ni, approximate ) values for these alloys can be computed, and a rough estimate of 7 for Cu(50 at. pct Ni) obtained by extrapolation. The relatively low y value obtained in this way for Cu(50 at. pct Ni) coincides with the observation of Nakajima19 that the stacking fault probability is a maximum at this composition. Likewise, the y values for the three ~u(~i) alloys are estimates based on values given by Swann and Nuttingo corrected on the basis Of Brown's estimate that the true 7 values are probably 2.3 times greater than those originally computed. In addition to the above alloys it was decided to investigate short-range ordered (s.R.O.) CU(AU) alloys containing up to 25 at. pct Au. This last alloy has been shown to have a well-defined planar arrangement of dislocations when deformed,21,22 probably due to cross slip being restricted by the S.R.O. Engineering fatigue data was to be represented by the determination of endurance strengths, and these were to be correlated with fatigue hardening and mechanical property data. EXPERIMENTAL I, order to study the desired properties and property changes, the following alloys were prepared: Cu(2.5 at. pct Si), Cu(5.0 at. pct Si), Cu(7.5 at. pct Si), Cu(5.0 at. pct Ni), Cu(25.0 at. pct Nil, Cu(50.0 at. pct Ni), Cu(5.0 at. pct AU), Cu(15.0 at. pct AU), Cu(25.0 at. pct AU). The copper and gold were zone-refined, while the silicon was semiconductor grade. The nickel was of 99.95 p,t. purity.* All alloys were induction-melted in a *Kindly supplied by the International Nickel co. helium atmosphere, appropriate precautions being taken to avoid segregation in the Cu-Au series. Rod stock was obtained by rolling and swag,ng. A few Bridgman single crystals were grown for fatigue-hardening studies, but most material was machined into polycrystalline fatigue speciments of the design shown in Fig. 1, and into polycrystalline tensile and fatigue-hardening specimens of simple cylindrical design. All specimens except Cu(Au) were annealed; the latter were quenched from above T, to produce short-range order. A few of the quenched alloys were examined for long-range order by step-scanning over

Jan 1, 1969

By F. B. Sherman, Keros Cartwright

Sanitary land filling has become one of the most widely used methods of disposing of solid refuse. A principal concern of regulatory agencies and the public itself is that landfill operations do not degrade the physical environment, including water resources, and the ground-water reservoir in particular. Knowledge of ground-water and engineering geology can guide landfill operations into suitable terranes or develop measures to compensate for natural limitations at a particular site. Experience and research in Illinois suggest four activities relating to landfill disposal that warrant attention by geologists and engineers: (1) regional delineation of favorable and unfavorable hydrogeologic conditions to facilitate planning and preliminary screening of potential landfill sites; (2) site evaluations, with considerations of geologic materials, topography, water levels, flow systems, and local occurrence and use of water resources; (3) research on aspects of the hydrogeologic environment that control effects of, or are modified by, landfills; and (4) formulation of practices in the siting, construction, and operation of landfills that prevent, mitigate, or isolate deleterious effects. The first two activities are basically in the domain of earth science, requiring the application of fundamental concepts of geology and hydrology and conventional site-exploration methods. The third activity, research, requires contributions from geology as well as other disciplines, including soil physics, sanitary engineering, and chemistry. The fourth calls for policy decisions by regulatory agencies and elected officials, using the contributions of scientists and engineers. Throughout history, man has disposed of unwanted materials by dumping. As urbanization has increased, haphazard dumping practices have given way to disposal under more controlled conditions because of increasing congestion of population and production of waste and greater concern for public health and environmental amenities. Many states and communities have already outlawed open dumping and open burning of refuse. The only practical methods of disposal of large volumes of refuse, therefore, are contained, high-temperature incineration, or burial in a sanitary landfill. In Illinois, regulation of solid waste disposal has been delegated to the Environmental Protection Agency. Each session of the legislature since 1965 has passed increasingly strict laws regulating waste diposal. As a result, the work of evaluating sanitary landfill sites has increased significantly for both the Department of Public Health, now the Environmental Protection Agency, and the Illinois State Geological Survey, which advises the Agency on matters of ground-water geology and pollution. In fact, our ground-water staff spends as much time on studies relating to waste disposal, primarily sanitary landfills, as on ground-water resource studies. Many other geological agencies are experiencing similar demands for increased assistance in solving waste-disposal problems. This paper summarizes some of the salient features of the sanitary landfill concept, describes activities of the Illinois Geological Survey in ground-water and engineering geology relating to landfills, and suggests policies that need consideration. A sanitary landfill is located and operated in such a way that vermin and pests, nuisances, and degradation of air and water are kept at acceptable levels. Some of the physical requirements of a sanitary landfill are all-weather roads for year-round access, fences to retain blowing paper, a daily cover of at least six inches of suitable earth material, and a final cover of at least 2 ft of earth material. Dumping into or adjacent to standing water generally is not allowed. Two common operating techniques are used. In the first, trenches are dug, the refuse is placed in them, and the earth removed from the trenches is used to cover the waste. In the second method, area fill, refuse is placed in low ground and covered with earth from adjacent high areas. The hydrology of the site is a prime consideration in locating sanitary landfills. Putrescible refuse, if saturated above field capacity, produces a leachate that usually has a high concentration of dissolved solids.2 As the leachate also acts as an agent for transporting bacterial pollutants, it constitutes a potential pollution hazard. To reduce the production of leachates, the topography of the landfill area should be such that surface water will not flow into or through the fill. Operations that will result in refuse disposal below or near the highest known water-table elevation may be required to take corrective or preventive measures to protect the ground water. In practice, under humid conditions such as prevail in Illinois, locations where disposal can take place above the water table are relatively few because surficial materials are commonly fine-grained, which permits slow gravity drainage and results in high degrees of saturation (100% moisture content) near the surface. At some sites, although disposal has taken place above the water table, ground-water mounds have developed, resulting in permanent saturation of the refuse. Permeability barriers usually are required to protect the ground-water reservoir from degradation by leachates. The convention in Illinois is to have a minimum of

Jan 1, 1972

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, 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 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 AIME AIME

A MODERN copper reduction works has both a genera1 chemical laboratory for control work and a research laboratory for the study of improvement of present processes and better working-up of by-products. The division of work in the general laboratory would naturally differ in different plants. The following statement to the writer from a chemist in one of the works will give a good idea of the type. of work in the general laboratory :

Jan 1, 1929

By Eleanor Sonley

In October 2000, temporary seismographic stations were installed above the Trail Mountain Mine, Emery County, Utah, USA to supplement a regional network. Over seven months, 1826 mining-induced seismic events were recorded with a magnitude range of M 0 to M 2.2 (Arabasz et al., 2002, 2005). Routine event locations for this dataset cluster around the longwall mining activity during the recording interval. However, applying a double difference relocation technique to the routine locations significantly improves the correlation between mining activity and event location. The resulting relocations depend greatly on station distribution and the number of arrival time picks. We use this dense temporary array to assess station distribution requirements for applying the double difference technique at other mines by systematically removing data from available stations, thus simulating sparser networks. The insight gained using the Trail Mountain dataset is used to develop a metric for assessing the usefulness of above-mine networks and to develop a tool for expanding seismic network coverage. To verify the appropriateness of these tools for wider application, we examine data from two additional mines located in the western United States. Improving the relative locations of mining-induced seismic events highlights the association of these events with active portions of the mine. Clusters, regions of quiescence and relative timing of events can be used by mine operators to assess hazard, optimize production, and evaluate performance of mine plans. The lessons learned from this study may be used in future network design to improve accuracy of event locations at other mines.

Jan 1, 2009

By Ali Kahriman, Alper Celtikc, Serdar Celik, Tuncer Gungor

This paper presents the results of ground vibration measurements carried out in Gülyali Hisarcik Basalt open pit mine located on the north side of the Anatolia near Ordu province in Turkey. Within the scope of this study in order to predict peak particle velocity level for this site, ground vibration components were measured for 5 shots during bench blasting. In blasting operations; ANFO (blasting agent), gelatin dynamite (priming) and delay electric detonators (firing) were used as explosives. Parameters of scaled distance (charge quantity per delay and the distance between the source and the station) were recorded carefully and the ground vibration components were measured for all blast events using 5 different vibration monitors (4 White Mini-Seis and 1 Instantel Minimate Plus Models). The absolute distances between shot points and monitor stations were determined using GPS. The equation of square root scaled distance extensively used in the literature was taken into consideration for the prediction of peak particle velocity.

Jan 1, 2013

By Russel R. Bohn

The objectives of this study were to: (a) determine the state of the art for the control effectiveness of various chemical dust suppressants, (b) field test the effectiveness of the dust suppressants, and (c) develop technical guidelines for selecting and implementing cost effective tailings pond dust control programs. Chemical stabilizers evaluated under field conditions were: Coherax, Calcium Lignosulfonate, Nalco 656, 655, and Magnesium Chloride. Testing in the field utilized a portable field in-situ wind tunnel. The wind tunnel was placed on the tailings surface and simulated wind speeds of up to 80 mph were used to determine wind erosion threshold velocities. Emission rate testing was performed with the wind tunnel through the use of a real-time cascade impactor. A control cost effectiveness evaluation was performed to determine how efficient the chemical stabilizers were on a dollars expended per acre basis.

Jan 1, 1983

By D. C. Moore

The decision to proceed on a "fast track" construction project is not" without risks as well as rewards. An understanding of some of the risks prior to start of the project will maximize the rewards. This paper describes the research, development, procurement, engineering and construction of the Asamera Minerals cannon Mine concentrator. This fast track project was accomplished in 20 months. Special attention is paid to the risks encountered and the steps taken to ameliorate them.

Jan 1, 1988

By M. E. Lewellyn

A new class of thionocarbamates, O-alkyl-N-allyl thionocarbamate (III), has been found to give better metallurgy than 0-isopropyl-N-ethyl thionocarbamate (I) for a number of sulfide ores containing copper, lead, and zinc. Data on the results of flotation tests with several such ores are presented. The improved metallurgy is believed to be due to the participation of the allyl double bond in the complex formation of the thionocarbamate on the surface of the mineral.

Jan 1, 1985

By Stephen Phillips, Luigi P. Gonano

Vertical circular shafts have been proposed for multiple access to high level nuclear waste deep geologic repositories. Shaft construction is a critical path item in the commissioning of a repository. Other constraints on the shaft design and construction criteria include the lack of design precedence for long term performance and the importance of minimizing rock and groundwater disturbance. A comparative evaluation of conventional and mechanized shaft sinking techniques is presented in the context of the particular geologic and design constraints. Technical and non-technical factors for blind drilling, blind boring, raise-reaming, drill-and-blast and ream-and slash methods are reviewed for typically "hard rock" and "salt" type repository sites. Appropriate shaft designs are developed, the suitability of each method to the particular requirements evaluated and construction costs and schedules determined. The explicit appraisal of the significant design factors for each method serves as a rational basis for the preliminary selection of the most suitable sinking technique for a particular set of site/design conditions.

Jan 1, 1983

By Geoffrey Purcell

The flotation response to amine collector of quartz and the feldspars microcline and albite, of the pyroxenes augite and spodumene, and of olivine can be, explained by a physical adsorption model. Flotation decreases in the order given above, and the different behavior of the three groups is attributed to structural variations of the silicates, With sulfonate collector both chemical action and physical adsorption play their part in producing hydrophobic surfaces, and flotation increases in the order given.

Jan 1, 1966

By R. Sullivan, K. Joiner, B. Emery, M. de Vries, F. A. Stober, L. R. Nelson, J. M. A. Geldenhuis

Keywords: Pyrometallurgy, smelting, converting, furnace design, furnace retrofit, high-intensity smelting, copper coolers, plate coolers, waffle coolers, tap-blocks, lower sidewall air-cooling, wall hold-down system, platinum group metals, ferroalloys, ferrochrome, ferrocobalt, ilmenite smelting, ironmaking, operational readiness, operational support The paper describes the development of furnace designs by Hatch in conjunction with the smelting plants in Africa to meet the intense process requirements in certain applications; continued improvement in operating efficiency through increased throughput from existing crucibles; and improvement in campaign life and furnace integrity. The era of Hatch in Africa has seen the doubling of furnace power in retrofit projects using existing crucibles to developing the highest intensity immersed electrode operations in the world. This has resulted in minimized OPEX and CAPEX per unit of production. Through the continued development of its cooling, binding and furnace power supply technologies and working with the experienced and knowledgeable personnel at the smelting facilities in Africa, Hatch has managed to meet the challenges of ever increasing furnace process requirements associated with increased power density and superheats prevalent in the operations. In addition to developing furnace crucible designs, Hatch has also intensified its ?after sales service and support? with the construction, commissioning and start-up technical assistance and operational readiness and operational support for ramp-up to nameplate capacity and beyond. The key areas of furnace risk associated with high superheat molten material tapping has also seen the development of diagnostic systems to mitigate risks and produce early warning signals for the operators.

Jan 1, 2006

By Brad Murray

Evaluating the performance of a shielded tunnel boring machine (TBM) can be difficult because the ground cannot be observed directly and in relation to the TBM attitude within the excavation. It is difficult to determine whether performance problems are caused by ground loading, steering issues, fine muck packing against the shield, or migration of annular grout around the shield. The Arrowhead Tunnels Project in Southern California uses strain gauges on its shielded hard rock TBMs to detect shield strains early and correlate them with tunneling conditions, thereby increasing the chances that problems will be correctly diagnosed and addressed to optimize TBM performance. The frequent automatic measurement of shield strain has also been valuable by confirming whether remedial methods used to free the machine after stoppages are effective. This paper discusses the qualitative interpretation of characteristic strain patterns on the Arrowhead Tunnels TBM shields and discusses installation details and special considerations for setting up a strain measurement system.

Jan 1, 2008

By S. Holtermann

Through the process of Riverbank Filtration (RBF), the Louisville Water Company plans to obtain a higher quality source of water for their B.E. Payne Water Treatment Plant (WTP) by means of four collector wells located in the alluvial aquifer along the bank of the Ohio River. The collector wells are connected to a 2,850 L/s (65 mgd) pumping station by a 2,362 m (7,750 ft) long, 3 m (10 ft) finished diameter tunnel located in limestone and shale bedrock, which is approximately 45 m (150 ft) below the surface. High tunnel boring machine (TBM) production rates were achieved in this moderately weak bedrock; most of the tunnel only required support in the crown. An unanticipated water-bearing fault and unanticipated ground behavior proved challenging and required quick thinking to minimize impacts to construction. Tunneling conditions are discussed in detail, together with the methods used for overcoming the groundwater and ground behavior challenges. The tunnel is expected to be completed in April 2009.

By Kyle Ott, Bernard Daly, Syed Haq, Douglas Anderson, Parsons Brinckerhoff

Drill and blast was determined to be the most cost-effective means of excavating the Raw Water Tunnel (RWT) for the Croton Water Treatment Plant in the Bronx, New York City. One of the major issues confronting project designers was the potential impact of blast vibrations on a major City Distribution Tunnel (CDT). The proposed path of the RWT would cross less than 30 meters (100 feet) above the CDT, which was built in 1917. Discussions with New York City Department of Environmental Protection led to vibration criteria on the CDT of 51 mm/sec (2 in/sec) peak particle velocity as an alert level, and 38 mm/sec (1.5 in/sec) as a review level.

Jan 1, 2012