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By W. H. Ode, W. A. Selvig

In 1938 , the Bureau of Mines cooperated with Subcommittee XV on Plasticity and Swelling of Coal ( C. C. Russell , chairman ) , of Committee D-5 on Coal and Coke , A.S.T.M. , by testing various methods for determining expansion and plastic properties of a series of coals of widely different coking properties . One of the tests made by the Bureau was the British Standard Method for the Crucible Swelling Test for Coal.5/ Briefly , this swelling test consists in heating 1 gram of pulverized coal in a silica crucible over a gas flame under prescribed conditions to form a coke button which is compared in size and shape with a series of standard profiles numbered 1 to 9. These profiles are intended to cover the size and shape of buttons obtained from coals covering a wide range in plastic properties .. The number of the standard profile with which the maximum cross -sectional area of the button most nearly corresponds is recorded as the swelling index . Although the swelling test showed poor relationships with the various tests that were used for measuring expansion or contraction of coal during carbonization , it gave indications of having value in measuring free - swelling and caking properties of coals when burned on fuel beds . An opportunity to investigate this occurred when an engineer of an eastern public utility 'company reported pronounced differences in burning characteristics on grates of various shipments of Pocahontas coal . Some of these coals gave trouble by swelling 'excessively and clogging the grates . Samples were ob- · tained , and it was found that the troublesome coals gave decidedly higher free -swelling numbers than did the coals that performed satisfactorily . Also , it was found that the free - swelling index test gave better differentiation between coals than could be obtained from the coke buttons from the · ´`standard volatile -matter test or from agglutinating - value results . The results of the free - swelling tests were considered by the Subcommittee on Plasticity and Swelling , and a special group of the subcommittee , under the chairmanship of R. A. Sherman , was delegated to study the test method . Ostborg , Limbacher , and Sherman , ofthe Battelle Memorial Institute , ' investigated the various factors affecting the test and found that the swelling index was a good indication of the suitability of coal for performance on residential -type stokers from the standpoint of the coking characteristics of the coal . Data were presented also showing that the indexes wer helpful in the selection of coal for industrial underfeed stokers .

Oct 1, 1946

By I. L. Sills

Introduction So many minerals and mineral products are used in the glass industry that it would require an unduly long paper to deal adequately with all of them. Only those that go to make up the batch for the more common types of glass will be discussed here, and even with these only, the list is a lengthy one. It is true that for the body of the glass comparatively few materials are required-as silica sand, soda-ash, lime, and feldspar. The others are used mainly to colour or decolorize the glass, or to give it some other desired characteristic, such as high lustre, greater physical strength, or opalescence; but, although they may be present in very small quantities, they are just as in-dispensable to the glass industry as are the main constituents of the batch. Sand Sand is essentially quartz, the dioxide of silicon, Si02. Most glass contains from 71J to 75 per cent silica. Sand for use in the glass industry, or 'glass sand', is obtained both from deposits of unconsolidated material, commonly referred to as 'sand', and from solid rock-sandstone, quartz, or quartzite. For the manufacture of glass, the material must possess a high degree of purity, as noted below, so that, although sand is widespread in most parts of the world, and sandstone, quartz, and quartzite are not at all uncommon, it is only in certain places that suitable deposits are found. At the present time, most of the glass sand used in Canada is imported from France, Belgium, and the United States. The French material, from Fontainebleau, near Paris, is known the world over as a very high quality glass sand. That from Belgium is nearly as good and is lower is price.

Jan 1, 1934

By T. Bhambhani

Research at Cytec over the past two decades has shown that the traditional approach for ore metallurgical characterization (typically conducted via laboratory-scale batch flotation tests), which involves changing reagents and conditions one-factor-at-a-time, leads to several pitfalls. The chemical solution thus selected has a narrow un-optimized window of performance and will likely fail in the plant. A holistic approach, called the Flotation Matrix 100? process for ore metallurgical characterization, was developed at Cytec about a decade ago and has been successfully used to develop robust reagent solutions for several operations around the world. The premise for the approach is that flotation outcome from ores is variable and is determined by complex interactions between various physical, chemical and operational factors. A key component of the process is the use of efficient multivariate designs focused on the chemistry aspects (reagents and pulp chemistry), which helps identify the best processing conditions for each ore type, as well as the conditions resulting in poor performance. It also includes advanced statistical methods for dealing with the complexity and variability of experimental data. The holistic approach is illustrated in this paper with a case study on Pb/Zn ores. Using just 18 carefully designed batch tests, it is possible to partially simulate the behavior of each ore type in the plant. The best conditions for each ore type, along with the conditions to avoid, are identified. Along with this, insights are obtained into the reasons for high reagent dosages for certain ore types, and interactions between variables. Keywords: flotation, holistic approach, metallurgical characterization, multivariate experimental designs, base and precious metals recovery, sulfide collectors and modifiers, Pb/Zn sequential flotation

Sep 1, 2012

By H. Y. Sohn, M. Y. Mohassab-Ahmed

"As part of research to develop a novel flash ironmaking process using reduction of iron ore concentrate by natural gas or hydrogen, sulfur and phosphorus distribution ratios between MgOsaturated slag and iron were determined in the temperature range 1550-1650°C under three different gas atmospheres: H2/H2O, H2/H2O/CO/CO2, and CO/CO2. The distribution coefficients of sulfur (Ls) were 4.0, 2.0, and 0.8, respectively, under these atmospheres. The distribution coefficients of phosphorus (Lp) were, respectively, 8, 17, and 20 under these atmospheres. In the flash process over -25 wt% P and S volatilize during ore concentrate downward flight in the reactor shaft in addition to avoiding coke. Therefore, iron produced by the flash ironmaking process would have much lower sulfur contents than the typical hot metal produced by the BF.IntroductionGiven the issues related to large carbon dioxide emissions and energy consumption in the existing ironmaking technologies, a novel flash ironmaking technology was conceived by Sohn [1-3] at the University of Utah. This technology is the first ironmaking process that will convert iron ore concentrate directly to metallic iron in-flight. This process will produce iron directly from concentrate without requiring pelletization or sintering and avoid the need for coke. In addition, this process concept takes full advantage of the fine particle size of concentrate with a large surface area, which permits rapid reduction by a gas. In this novel ironmaking process, potential reductants and fuels are H2, natural gas, and coal gas. Therefore, an iron-slag bath is expected to be at equilibrium with gas atmospheres of mainly H2/H2O or CO/CO2/H2/H2O, respectively, corresponding to the use of H2 or natural gas/coal gas. Since the effects of an H2O- containing atmosphere on slag properties have not been studied before, it is critical for the development of the novel flash process to determine these effects. In H2-H2O-CO-CO2 gas mixtures under ironmaking conditions, H20 was found to have the most pronounced effect on the slag chemistry [ 4, 5]. In addition, it was found in this laboratory that water in the gas atmosphere affected the distribution of Mn between slag and molten iron [6]. Most previous work used CO/CO2 gas mixture to control oxygen partial pressure (p02) [7, 8]. Others used only inert atmospheres [9], although Kor and Richardson [10] used a mixture of C02/H2 to control p02. For phosphorus, previous research focused on Ar [11], Ar-CO mixture [12], CO [13], CO/CO2 [14] or N2/CO [15]. Therefore, it was necessary to investigate the effect of water vapor content in H2-H2O-CO-CO2 mixtures on the equilibrium distribution of sulfur and phosphorus between CaO-MgOsatSiO2-Al2O3-FeO- MnO (0.2-0.8 wt%)-P2O5 (0.1-0.9 wt%) slag and molten iron as part of the development of the novel flash ironmaking technology."

Jan 1, 2014

By R. K. Clifford

Chalcopyrite (CuFeSz) concentrates produced on the Lead Belt of Missouri typically contain 28 pct Cu and from 0.5 to 1.5 pct each of nickel and cobalt. Most of the nickel and cobalt are present as the mineral siegenite, (Ni,Co)3S4, almost none of which is recovered presently. The Bureau of Mines, U.S. Department of the Interior, has been conducting research on the chalcopyrite to produce a marketable fraction enriched in cobalt and nickel. Ferric chloride leaching of various chalcopyrite (copper) concentrates has yielded residues containing from 10.2 to 16.2 pct Ni and 8.3 to 11.4 pct Co with recoveries of approximately 79 pct. In another approach, laboratory froth-flotation tests have been conducted to investigate the effects of regrinding, roasting, baking, and steaming prior to flotation, with a variety of reagents. The best laboratory results were obtained after regrinding the copper concentrate to 94 pct minus 400 mesh and floating the chalcopyrite at natural pH (6.5) with 0.48 pound of sodium diethyldithiophosphate and 0.01 pound of sodium cyanide per ton of concentrate. This procedure yielded a tailings assaying 20.4 pct Cu, 5.2 pct Ni, and 3.45 pct Co. The tailings contained 67.2 pct of the nickel and cobalt and 15.9 pct of the copper originally in the copper concentrate; the remainder of the values reported to the flotation concentrate, which analyzed 29.6 pct Cu. To verify the laboratory flotation results, a 60 kg/hr process demonstration unit (PDU) was constructed in the Cominco American Magmont mill, Bixby, Mo.

Jan 1, 1978

By F. M. Makarinsky

"SynopsisThe Kanichee open pit mine and milling complex is situated approximately four air miles north-west of Ternagami, Northern Ontario, in Strathy Township. Mining and milling production commenced in January 1974.Milling throughput currently is at a rated capacity of 500 tons per day, seven days per week. A bulk nickel-copper sulphide concentrate is produced, and is trucked daily to the smelter at Falconbridge Nickel Mines Limited, Falconbridge, Ontario, some 120 road miles from Kanichee. Concentrate grades produced are in the order of 4.0% Ni and 8.0% Cu, with corresponding recoveries of 68-70% nickel and 86-90% copper. Stage additions of 100% guar-gurn solution reagent to the flotation cells for depression of talcose minerals, proved to be the major break-through for optimizing metallurgical performance.2. HistoryThe present Kanichee Mining property is known to old-timers in the Temagami area as the former Cuniptau Mine. Several attempts were made in the mid '30's and 40's to extract the ore at a profit, but each effort resulted in failure.In 1933-36, Cuniptau Mines Development Co. Limited, later called Cuniptau Mines Limited, sank a 245-foot vertical shaft and completed 2,200 feet of lateral work on the 100-140 and the 225 foot levels. Fifteen tons of ore were sent for testwork to the Mines Branch, Ottawa, Ontario.Cuniptau Mines Limited, in 1936, milled some 3,000 tons, and the resulting concentrate was put through a pilot smelter situated on the present mine site. Metallurgically, the smelting results were not encouraging, primarily due to the high melting point of the talcose minerals in the concentrate. From the period 1937-48, minor exploration work was done by Ontario Mining Corporation, Ontario Nickel Mines Limited, and Trebor Mines Limited. During the period 1948-49, 128 diamond drill holes, totalling 40,900 feet were evaluated by Trebor Mines Limited. In 1961, the name was changed to Ajax Minerals Limited, during 1968, five diamond drill holes were sunk, totalling 1,700 feet. Since that period the property has remained idle."

Jan 1, 1975

By N Robb

This paper gives an account of the laboratory testing, plant testing and final inclusion of the two-stage sulphuric acid leaching process in the Milliken mill. Operational results for September, 1962, are presented, showing a uranium extraction of 95.5 per cent, with a saving of some 16 lbs. of 100 per cent H~S04 and some 9 lbs. of lime per ton milled. Methods for effecting further improvements in the process are indicated. HISTORY P RIOR to the incorporation of two-stage leaching, the process for the treatment of uranium ores at Milliken consisted of a single-stage sulphuric acid leach, followed by a solution recovery step involving washing in thickeners and two-stage filtration. The thickener overflow solution was subsequently concentrated and purified by ion exchange, and the uranium was recovered, by two-stage precipitation, as ammonium diuranate. Figure I illustrates this flowsheet. For satisfactory extraction of uranium from the. Milliken ore, a high residual acidity, of the order of 40 grams per litre of free acid in the discharge solution, must be

Jan 1, 1963

President MRS. HARRISON SOUDER Smith Paramus Road Ridgewood, N. J. First Vice-president MRS. ROBERT HURSH 10 Kilburn Road Belmont, Mass. Second Vice-president MRS. ERLING L. JORGENSEN 95 Knollwood Rd. Short Hills, N. .1. Third Vice-president MRS. PERCY E. BARBOUR 1R(; Riverside Drive New York, N. Y. Fourth Vice-president MRS. C. B. MURRAY 2040 Stearns Rd. Cleveland, Ohio

Jan 1, 1938

By William B. Hughes

The rubber-sleeve core barrel was developed to improve core recovery from unconsolidated sands, where it is most difficult to obtain cores with conventional barrels. The use of a rubber-sleeve core retainer, together with other departures from clsual design, has required a considerable amount of testing to prove its op-erability. Tests run in shallow experimental holes and in drilling wells in the U. S. and Vemzuela have, to a large extent, proved the operation of the barrel. Internal mechanism troubles, such as sandbinding of the sleeve, have been largely overcome. Cores have been recovered from completely unconsolidated sands where conventional coring has almost been abandoned because of poor recovery. In addition to retaining the core until brought to the surface, the barrel gives a "packaged-ar-cut" core, which is coilverzient for handling and transportatiorz. In the initial work, drag blade cutters, roller cutters, and tungsten carbide insert cutters were usrd for cutting unconsolidated sands. Further bit head development is presently being conducted using diamond heads in unconsolidated sands where shaler, hard sands, and limestones are encountered in adjacent formations. Thir work should render the barrel capable of cratting and recovering most types of formations, possibly including broken and frac- tured fornlations. This development is being carried on in cooperation with a diamond bit manufacturer. INTRODUCTIO N For many years the oil industry has needed greater recovery of cores from unconsolidated formations. Conventional equipment has been used to core unconsolidated sands with little success. Specifically, a core barrel was needed which would recover unconsolidated sands in a condition that would give useful reservoir information. DEVELOPMENT OF 'THE RUBBER-SLEEVE CORE BARREL Basic Requirements of Unconsolidated Sand Corino The basic need was defined as a core barrel that would obtain a continuous unconsolidated core with sand grains in place as deposited in the formation. A preliminary survey of coring equipment and search of the literature pointed out the weaknesses of conventional core barrels. Laboratory tests using unconsolidated sand in a metal tube simulated actual sand movement into the core barrel during conventional coring. In conventional barrels the core sometimes crumbles and bridges, or fails under compressive column action, becomes oversize, and wedges against the walls of the inner barrel. The core barrel requirements established by the surveys and preliminary tests are as follows. 1. The core should be held to- gether as it is cut so it cannot crumble and bridge against the walls of the coring tool, care being taken not to permit damage from fluid circulation. 2. The core should be continuously supported as it is cut so it will not have to support itself in compression. 3. The core should be packaged in the barrel to retain an undisturbed condition during the trip out of the well. The packaging should be such that the core can be removed from the barrel and shipped to the core analysis laboratory without physical disturbance or alteration of fluid content by exposure to the atmosphere. Operation of the Rubber-Sleeve Core Barrel The rubber-sleeve core barrel (Fig. 1) meets the basic requirements for the recovery of unconsolidated sands.

By Timothy C. Siegel

Given that piles are composed of different materials and installed using a variety of methods, ASTM D 1143/D 1143M-07 Standard Test Methods for Deep Foundations Under Static Axial Compression is justifiably general so that it has as many applications as reasonably possible. There are aspects of the test setup, the test procedures, and data interpretation for instrumented piles (some of which are specific to augered cast-in-place piles) that are either not discussed in detail in this ASTM standard or are not addressed at all. Specifically, the use of variable time hold times can obscure the shape of the conventional load-deflection curve, as well as, influence the interpretation of the ultimate pile capacity. Unload-reload cycles induce additional non-uniform internal axial loads within the pile which complicate interpretation of the strain gage data and increase the potential for errors. Attempts to maintain a constant top load during the hold period can lead to operational difficulties in taking measurements and are technically unnecessary. Residual load has been shown to develop in cast-in-place piles and can significantly influence the axial load distribution as interpreted from strain gage data. The stiffness of cast-in-place piles has been observed to vary with the measured strain, and a constant modulus, as provided by correlation with the unconfined compressive strength of the grout, may lead to a significantly different interpretation of the axial load distribution.

Jan 1, 2010

By Luther J. Russell

Mineral development in South Dakota is aggressively encouraged by both state statute and county ordinances. Yet, an increasingly conservative philosophy is taking hold, resulting in expanded regulatory controls and concomitant increases in direct and indirect costs to permit and operate a mining operation. This paper examines the hidden costs associated with protracted permitting timeframes, land buy-outs, attorney's fees, etc., associated with the increasing emphasis for stiff controls on mineral development in South Dakota.

Jan 1, 1998

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 * Pebble is a commercial designation for the coarser fraction of finished phosphate from a washer, usually +14 mesh. 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 * Pebble is a commercial designation for the coarser fraction of finished phosphate from a washer, usually +14 mesh. 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 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 B Besa

Australia is a world leader in the design and operation of mines accessed by the decline method. Literature indicates that a number of metalliferous underground mines using the decline system are also increasing steadily. This increase has generated a great deal of interest in underground haulage systems in future due to the huge costs involved in the development of declines. Studies have also shown that the current method of accessing orebodies by the conventional decline method has proved to be expensive, especially for narrow vein deposits. This is because decline development costs are influenced by mine planning and design parameters as well as available haulage systems that the mining engineer can choose from. The current access method suggests that more waste material is being extracted because of the size of decline openings adopted (5.5 mW and 5.5 mH). Also the conventional 1:7 decline gradient and turning radius of 20 m makes decline lengths, as well as the duration of development, excessively longer. This paper discusses results of a mine design case study with monorail technology as a method of accessing and developing the decline. The proposed system is economical, has high advance rates and offers a significant reduction in capital expenditure. The study compares costs, development metres and the duration of decline development using the monorail system with the conventional (truck haulage) access method. Results obtained indicate that the monorail technology has the potential to reduce the decline length by over 62.6 per cent and decline costs by 63 per cent for the case study. The monorail system also reduces the total capital development costs by 22 per cent for the same case study. Due to rapid development resulting from increased advance rates and the shorter decline length coupled with smaller decline openings, the time to access the deposit reduced by 71.8 per cent. This case study demonstrates that the economics of narrow vein deposits can be improved significantly by using the monorail system of decline development.

Jan 1, 2008

By A S. Fogliata, L Bengochea, G Mas, J E. Lazarte

The Culampajß auriferous district is located in the northern sector of the eastern edge of the homonymous hill, Catamarca Province, Argentina. Tajo Largo mine is situated in the southern flank of Las Minas gulch at an approximate height of 3.600 m above sea level. In the studied area main outcrops are orthogneiss rocks of Chango Real Formation. The ore mineral is native gold with a lesser amount of pyrite, chalcopyrite, haematite, limonite, malachite, azurite and covellite. These ore mineral assemblages are disseminated in quartz veins filling tension fractures in the host orthogneiss.Several types of inclusions have been recognised, including three phase H2O-CO2liq-CO2vap inclusions, polyphase inclusions with H2O-CO2 and one or more solid phases, and two phase aqueous inclusions. In the three phase H2O-CO2liq-CO2vap inclusions the Tmclath is about 1¦C and the CO2 liquid-vapour homogenisation occurs at 26 - 28¦C, to liquid phase. The homogenisation temperature of H2O-CO2 occurs at 365 - 380¦C in the aqueous phase. This set of data defined a solution with a salinity of 14, 4 wt per cent NaCl equivalent and a density of 0.84 g/cm3. In the two phase aqueous inclusions the melting temperature (Tm) of -5.4¦C and the eutectic point (Te) in -38¦C indicates a salinity of about 7.5 wt per cent NaCl equivalent, with the probable presence of Ca2+. Density of fluid is about 0.83 g/cm3. Homogenisation temperature ranges between 260¦ and 280¦C. The pressure, temperature and salinity of fluids in the polyphase inclusions suggest an early fluid of high density and high salinity. The three-phase fluid inclusions may be the result of an unmixing-phase separation of post magmatic hydrothermal fluid due to pressure decrease along fractures. It is suggested that the deposit was formed during the late magmatic hydrothermal stage genetically linked to granitic rocks and can be comparable with the intrusion-related gold system of Baker (2003).

Jan 1, 2008

An accelerated and coordinated program of coal research is believed by many to be a necessary step in assuring coal's proper position in the expanding energy market. The President's Advisory Committee on Energy Supplies and Resources Policy, for example, made the following recommendation: We recognize that coal is a great national asset and endorse a cooperative study to determine what research and development could be undertaken. The coal industry and both Federal and State Governments should participate in this study and its costs. Representatives of the coal industry have agreed that such a study should be initiated. Accordingly, Felix E. Wormser, Assistant Secretary - Mineral Resources, United States Department of the Interior, requested the Bureau of Mines and Bituminous Coal Research, Inc., to conduct a survey of current research on bituminous coal. In addition to this survey, studies were made to determine research projects that would be of value to the coal industry.

Jan 1, 1956

By Y. A. Rocher, J. E. Dorn, L. A. Shepard

Creep activation energies for single aluminum crystals favorably oriented for shear by (010) [101] glide were detemined over the temperature range from 78" to 900°K. Observations of slip bands on the specimen surface were made in conjunction with the investigation. From 78" to 780°K, the activation energies obtained in this imestigation agreed closely with those previously found for creep by (111) [101] slip. Between 78" and 140°K, the activation energy was identified with the Peierls process, while between 260°and 780°K the activation energy was close to that for cross-slip. The coarse wavy slip bands nominally parallel to the (010) plane observed above 260°K were attributed to fine cross-slip. From 800" to 900°K, unusually high apparent activation energies ranging from 28,000 to 54,000 cal per mole were obtained. These apparent activation energies were attributed to re crystallization. AS illustrated in Fig. 1, a recent investigation1 has shown that creep of aluminum single crystals by the (111) [i01] mechanism is controlled by three unique processes, each of which is characterized by a single activation energy which is independent of the applied stress and the creep strain. A comparison of the observed activation energies with theoretically calculated values permits a fairly clear identification of the three operative creep processes. Below 450°K, where the activation energy for creep is 3,400 cal per mole, the deformation is controlled by the Peierls process, the activation energy for creep agreeing well with that calculated by seeger2 for the energy required to nucleate the motion of a dislocation loop against the atomic forces of the lattice. Between 590° and 750°K, the observed activation energy for creep of about 28,000 cal per mole agrees well with the energy necessary to induce cross-slip. Seeger and schoeck3 estimate that the activation energy is about 24,000 cal per mole whereas Friedel4 recently calculated this activation energy to be 28,000 cal per mole. Above 800°K the activation energy of 35,500 cal per mole that was observed for creep agrees well with that estimated for self-diffusion in aluminum.= In this range the operative rate-controlling slip process has been clearly identified as that arising from the climb of edge dislocations. The objective of this investigation is to ascertain whether a single crystal of aluminum favorably oriented for simple shear in the [loll direction on the (010) plane might exhibit uniquely different activation energies for creep from those obtained previously for (111) [101] slip. Whereas the exis- tence of such unique activation energies would constitute incontrover table evidence for new mechanisms of slip, the absence of any new activation energies might suggest that slip of aluminum is confined to the (111) [loll mechanism. Several factors prompted the selection of the (010) [101] orientation for study. First, there are more reported observations of (010) [loll slip than of any other nonoctahe-dral mechanism.8-10Secondly, Chalmers and Martius1l have concluded from considerations of the energies of dislocations that (010) slip is the second most favored mechanism in face-centered-cubic metals. Finally, favorable orientations for simple shear by the (010) [loll mechanism provide the least favored orientation for slip by the (111) [101] mechanism. EXPE-RIMENTAL PRO-CEDURE The high-purity aluminum stock, specimen preparation, shear fixture, extensometry, and experimental technique used in this investigation were the same as those previously reported.' Single-crystal spheres grown from the melt of 99.995 pct pure Al* were _ *The high-purity aluminum used in this investigation was graciously given by the Aluminum Company of America. oriented, carefully machined into dumbbell-shaped shear specimens, annealed, and chemically polished. The finished specimen had a central reduced section 0.190 in. wide and 0.590 in. in diam and 1/4-in. grip sections at both sides, 0.690 in. in diameter. The specimen was oriented in the stainless steel grips of the shear fixture with the (010) plane perpendicular to the dumbbell axis and the [loll direction parallel to the stress axis within 2 deg. Creep activation energies were calculated in the previously described manner1 from determinations of the instantaneous change in shear strain rate produced by an abrupt 15 to 20 deg increase or decrease in test temperature. If is the instantaneous strain rate at strain y and temperature T1, and ?2 the instantaneous rate at y and T2,

Jan 1, 1960

By M. Botz

For leach solutions containing primarily free cyanide, the silver nitrate titration method is a rapid and acceptably accurate means of analyzing for cyanide. However, in leach solutions containing elevated levels of copper, the titrimetric method significantly over-estimates the concentration of free cyanide and under-estimates the concentration of weak and dissociable (WAD) cyanide. The authors have developed a modified potentiometric titration method for determining WAD cyanide that overcomes issues with high copper levels in solution. The new MP-WAD method involves addition of a chemical ligand to the sample prior to silver nitrate titration to fully release cyanide bound to copper. Once released from copper, cyanide in solution is no longer affected by the presence of copper and titration results are quantitative rather than empirical. In multiple laboratory trials with synthetic solutions and complex actual leach solutions, accurate WAD cyanide measurements were achieved using the MP-WAD method. Analyses obtained with the new method were found to be within about ±10% of analyses obtained with the USEPA OIA-1677 WAD cyanide analytical method.

Feb 27, 2013