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By L. A. Harris

The formation of coke during coal liquefaction can eventually result in catastrophic failure of the conversion system. Coking typically proceeds by the formation of liquid crystals (mesophase) and the subsequent separation of these crystals from the plastic phase formed during heating of coal. The anisotropy of the mesophase makes it readily detectable in the optical microscope under polarized light. This study examines the solids formed during coal liquefaction processing by optical and electron microscopy. Observations on the relationships between the coke morphology, minerals and inert macerals are presented.

Jan 1, 1982

By W. Joseph Schlitt, Kenneth J. Richards, John D. Prater

Kennecott Minerals Company has developed a leach-hydrolysis route for treating flue dusts from copper smelters The flue dusts are acid leached in an oxygen atmosphere to solubilize copper and fix the arsenic, antimony, and bismuth as stable compounds Copper is recovered by cementation with Iron and the barren solution is recycled Molybdenum and other values may be recovered by modification of the basic flow sheet Results of bench scale and pilot plant tests are discussed in terms of the effect of operating parameters.

Jan 1, 1981

By F. W. Maclennan

The following is a list of members who died in 1929. It is compiled from reports to the Secretary's office. Biographical sketches published in Mining and Metallurgy are indicated in the last two columns. Issue Containing Year of Date of Biography Election Name Death 1929 Page 1922 Adams, Waldo Peck................................ June 2 September 435 1903 Argall, Joseph.................................... Nov. 1 1894 Austin, Leonard S................................. Oct. 30 December 580 1905 Bailed, Melbourne.........................Dec. 12 May* 280 1890 Barringer, Daniel M.............................. Nov. 30 February* 119 1895 Batchelleb, HENRr R.............................. Jan. 15 March 167 BINGHAM Charles E............................. 1923 Bibkett, E. H..................................... Apr. 23 August 394 1913 Bibbell, Robert W................................. Jan. 18 April 214 1926 Blackmar, Don C.................................. June 4 March* 189 1920 Bosworth, T. O.................................... Jan. 18 April 214 1894 Boucher, Arthur S..........................Jan. 6 May 257 1927 Brandenthaler, R. R.............................. Dec. 14 January* 64 1889 Bretherton, S. E............................... Oct. 4 December 581 1911 Brunton, F. K................ .. Aug. 17 1924 Brush, Charles F................................. June 15 September 436 1925 Cameron, Frank N........................May 16 July 349 1891 Chambers, Robert E.............................. May July 348 1908 Clark, J. Murray................................. Dec. 3 August* 407 1883 Clymer, edward T............................. Oct. 13 December 579 1914 Cofpey, George T................................. May 8 July 349 1901 Cooke, L. H....................................... Aug. 23 November 537 1915 Cousins, George W........................ June 30 February* 118 1886 Cowles, A. H...................................... Aug. 13 September 436 1892 Crane, Theron I............................. Nov. 2 December 579 1921 Crowell, Samuel B................................ Dec. 9 January* 64 1907 Curtis, Bracey.................................... Jan. 31 April 215 1895 de Batz, Rene........ ............... January March* 189 1928 Dines, Tyson S.................................... March May 257 1909 Dufourco, R. G............................. January November 537 1920 Ericsson, Eric A.................................. Apr. 4 April* 231 1920 Erlenborn, Willi....................... Nov. 10 March* 189 1888 Farish, John B.................................... Nov. 14 December 580 1891 Ferguson, Vincent........................... Mar. 8 May 257 1917 Finucane, T. R.................................... Mar. 22 February* 118 1921 Foch, Ferdinand................................... Mar. 20 1885 Foucar, Edouard I....................... Jan. 3 March 167 1880 Francklyn, Charles G............................. Jan. 11 April 214 1925 Friedl, Arthur J.............................. July 9 September 435 1894 Fuller, James W. Jr............................... April 4 1876 . Goodale, Charles W.............................. April 11 May 258 1912 Graybill, John..........................Dec. 11 June* 320 1887 Gritzner, F. A.................................. Mar. 26 July 394 1906 Gunther, C. Godfrey.............................. Deo. 26 February* 118 1900 Haggen, Edward A.............................. Apr. 22 October 488 1895 Hall, Benjamin M........................Nov. 19 March* 189 1913 Harper, Walter S................................. Feb. 15 August 394 1921 Hawley, William S................................ Apr. 20 June 298 1920 Herd, Walter ... . . . ..... Aug. 23 September 435 1889 Holt, Marmadtjke B............................... Dec. 26 February* 118 1928 Husk, Albert D .... ................June 9 December 580 1920 Hutchinson, S. Pemberton......................... Feb. 16 April 214 1926 Ivanoff, Constantine L............................ Mar. 7 June 298 1914 Jaeger, Frederick..... .............July 28 October 488 1891 Jennings, Robert E.......................Oct. 30 December 580 1922 Johnson, Homar L................................. Sept. 17 October 483 1927 Johnson, Homer H................................. Mar. 21 May 257 1918 Jones, Jesse L ..... ...............Jan. 22 April 214 1892 Jones, Thomas J................................... June 18 October 488 1883 JoOet, Cavalier H................................. Mar. 29 May 258 1881 Kennedy, John S . . . . ... ..........Apr. 22 August 394 1881 Kennedy,- Hugh.................................... May 23 September 435 1895 Kebvin, James H................................... September October 488 1926 Kkom, L. J........................................ Dec. 24 January* 64 1905 La Follette, Harvey M........................... Sept. 19 November 538 1920 MacDonald, Alexander............................ May August 394 1904 Marengo, Paolo................................... May 257 1925 Maxwell, Allison R............................... May 7 November 538

Jan 1, 1930

By K. Regl, J. Berneder

"AMAG investigated different aspects to improve the resistance against grain coarsening in aerospace alloys during recrystallization and was successful in finding several opportunities to minimize this effect. The studies show that one aspect to counteract grain coarsening is the chemical composition and subsequently all phases that are formed during homogenization. Starting with a fine grain (~ 50 µm) can lead to a final grain size of few mm. On the other hand with a coarse initial grain size (approximately 150 µm) and the right chemical composition the grain size can stay almost constant despite different deformation steps. Additional we investigated different sequences for the component production to eliminate the problem of grain coarsening as a coarse grain has a detrimental effect on product performance.INTRODUCTION In the aircraft industry, aluminium sheets are most commonly deformed in their soft state, followed by solution heat treatment and natural or artificial aging. If the preceding deformation is in a critical amount, during the recrystallization at solution heat treatment grain coarsening can occur. A subsequent deformation to this condition leads to an orange peel surface, which requires additional polishing. Next to this extra effort the coarse grain size produces a drop in the mechanical properties, often below specification limits. If there is no subsequent deformation, this effect is not visible and without destructive testing there is no possibility to detect the error. As mentioned before, this effect occurs during recrystallization, which is a process that is driven by the stored energy in the microstructure. The two most important aspects for an increase of the stored energy are the increase of the grain boundaries fraction (smaller grain size) and the increase of the dislocation density caused by deformation. The increase in dislocation density during the deformation arise from trapping of existing dislocations and from the generation of new dislocations. During recrystallization all grains are built completely new, so all dislocations are removed (Humphreys & Hatherly, 2004). There are three requirements for recrystallization:"

Jan 1, 2018

Electrostatically charged water droplets efficiently suppress breathable dust produced during open air rock crushing operations, according to recently completed independent tests conducted for the Environmental Protection Agency. Since a primary rock crusher produces very heavy clouds of dust, this application was selected as an appropriate test site to measure the efficiencies of new "charged fogging" equipment in mining and mineral processing operations. EPA's Industrial Environmental Research Laboratory, Research Triangle Park, NC, contracted TRC-Environmental Consultants Inc. of Wethersfield, CT, to conduct a full-scale demonstration of a recently developed electrostatic fogger, designed to project a charged fog spray at distances up to 15.2 m. Test results indicated overall fogger efficiencies of approximately 70%. The Fogger IV* employed in these tests is reportedly the largest charged logger ever made, with an extremely long operating range. The original fogger was designed for indoor dust control; this fourth generation version is designed for fugitive dusts, indoor and outdoor, and has an operating range up to 15.2 in.

Jan 6, 1981

The Modified Zincex® Process, developed by Técnicas Reunidas, is highly flexible dealing with a wide range of pregnant leach solutions. No matter if those solutions are produced from primary or secondary raw materials, and/or are based on conventional leaching, pressure leaching, heap leaching or bio leaching the result is always high-purity zinc product. The zinc solvent extraction unit within the process is a very effective barrier for impurities and a buffer for the composition of the pregnant liquor. This unit achieves an extremely pure aqueous extract that can be converted into different final products. All of them are characterized by fulfilling the highest purity standards. Also, there is the possibility to adapt the global process to a wide range of pregnant liquor compositions and various production capacities up to a zinc production limit over 150 000 t/a with a single solvent extraction plant.

Jan 1, 2002

By M. C. Udy, F. W. Boulger

AN incomplete phase diagram for the U-Ti systern was determined earlier 1 and more recently, a tentative diagram was presented for the uranium-rich end of the system.' In the present re-examination of the whole system of U-Ti alloys, high purity materials were used. Melting stock for the alloys was high purity uranium, containing about 0.09 pct C as the only appreciable impurity, and high purity iodide-process titanium purchased from New Jersey Zinc Co. Both metals were cold rolled to about 1/6 in. thickness, sheared to about I/' in. squares, and cleaned by pickling. The alloys were arc melted under a helium atmosphere in a water-cooled copper crucible. A thoriated-tungsten electrode was used. The furnace chamber was evacuated, then flushed with helium, prior to each melting. It was finally filled with stagnant helium at one atmosphere pressure. Each alloy was remelted three times after the original melting, to insure homogeneity. The alloy button was turned bottom side up before each re-melting operation. Some 22 alloys were examined. Their compositions were spaced at appropriate intervals between 100 pct Ti and 100 pct U. Analyses were made on chips taken after fabrication. The major contaminant was carbon, which varied from 0.03 to 0.08 pct. It appeared in the microstructure as titanium carbide. Alloy compositions were calculated to a carbon-free basis for consideration on the diagram. Tungsten and copper, possible contaminants from the melting operation, were generally less than 100 parts per million each. Fabrication All alloys were forged and rolled to bars approximately V8 in. square. They were clad either in SAE 1020 steel or in a 5 pct Cr-3 pct Al-Ti-base alloy, depending on the fabrication temperature. A temperature of 1800°F (980°C) was used for alloys near the compound composition. This necessitated using the titanium-base alloy, since iron reacts with titanium at this temperature, producing a low melting alloy. Other alloys were fabricated at 1450°F (790°C), using steel jackets. No iron-titanium reaction occurred at this temperature. The jackets were welded in place in an argon atmosphere. Those alloys sheathed in steel were declad and then reclad between rolling and forging operations. On the other hand, those clad with the titanium alloy were cut to a roughly rectangular shape prior to clading and were then carried through both the forging and rolling operations without opening. Those alloys near the compound composition were found to be cracked when the clading was removed. The cracked materials had been plastically deformed, however, and at least some of the cracking had OCcurred during cooling. Heat Treatment The rolled bars, after being declad and shaped to remove surface contamination, were all given an homogenizing treatment of 160 hr at 2000°F. (Samples were taken for analysis following the declading and shaping operations.) All were heat treated at the same time in one furnace, but each was sealed in a purified argon atmosphere in an individual Vycor glass tube. Argon pressure was such that it was approximately atmospheric at temperature. One end of each tube contained titanium chips and this end was heated to 1200°F (650°C) for 10 min prior to the heat treatment. This purged the atmosphere of residual reactive gases. The balance of the tube was warmed during the purge to liberate adsorbed moisture and gases, which also reacted with the hot chips. The bars were furnace cooled from the homogenization treatment. Specimens of each alloy were water quenched after 2 hr heating at 1000°, 1200°, 1400°, 1600°, 1800°, and 2000°F (540°, 650°, 760°, 870°, 980°, and 1095°C). In addition, some were treated at intermediate temperatures of 1300°, 1500°, and 1700°F (705", 815", and 925°C) and at 2150°F (1175°C). Specimens, about '/s in. cubes, were cut from the bars, sealed in individual Vycor tubes, and heat treated as described. All specimens heat treated at the same temperature were processed together. Samples were quenched by breaking the Vycor tube rapidly under water. Metallographic Examination Specimens were mounted in bakelite and ground wet on 180 grit paper held on a 1750 rpm disk. They were then ground wet by hand, using 240, 400, and 600 grit papers. The rough grinding was continued long enough to get well below the surface. Specimens were mounted separately because of the variation in the rate of etching between alloys. The specimens were polished with rouge on a 4 in., 1725 rpm wheel covered with Miracloth. Alloys on the titanium side of the compound composition were etched with a solution of 2 pct hydrofluoric acid in water saturated with oxalic acid. A few crystals of ferric nitrate were added as a bright -ener. Specimens were immersed 5 sec, polished to remove the etch, then re-etched. With the higher titanium alloys, it was often necessary to start the etch on the polishing wheel, because of the formation of a passive film. In some instances, a plain 2 pct hydrofluoric etch was satisfactory. For the alloys on the uranium side of the compound, a distinction between the compound and the uranium phase developed after standing a short time in air. This could be hastened by the application of heat, such as obtained by placing the specimen on a radiator. A deep etch was necessary to develop details in the uranium-rich phase, such as the Widmanstaetten pattern sometimes obtained by quenching y uranium. A 2 pct hydrofluoric acid solution was used for this deep etching.

Jan 1, 1955

By Jing Zhang, Qing Xiao, Xulong Liu, Qiuju Li

"Pickling sludge generated from stainless steel production process contains large quantity of Cr, Ni, and Fe. The disposal of and fill sludge directly not only leads to the heavy metal dissolves in the soil and underwater, but also loss value metals. However, the major mineral phases of value metal are hydroxides and there are lots of calcium fluoride and water in the sludge. So the pickling sludge has to be pretreated before reduction-magnetic separation. The TG analysis, reduction reaction and magnetic separation were all used to study the pickling sludge. The result indicated that up to 900°C the total weight loss of pickling sludge was 22.32%. The mass contents of Fe, Cr and Ni in the tailings (non-magnetic) were 4.27%, 0.067% and 0.041% respectively.IntroductionSince 21st century, the stainless steel production in the world has been continuously increasing because of its excellent performance [1]. However, high stainless steel production means a high pickling sludge production. For example, in a stainless steel plant with waste water quantity of 100m3 /h, the production of pickling sludge will exceed 15 Kt/a. Pickling is essential for generating of a homogeneous and corrosion-resistant surface during stainless steel production subsequent process. Pickling of stainless steels is predominantly carried out with an aggressive solution, sulfuric acid H2S04 or composed of nitric acid HN03 and hydrofluoric acid HF [2]. During the pickling process, the pickling acids are recovered through ion exchange and metal ions such as Fe3+, Cr3+ and Ni2+ get into the acids forming the pickling waste water [3]. Then, the pickling waste water is precipitated with lime, forming pickling hydroxide sludge. It is not feasible to recycle or treat the pickling sludge using conventional methods, because the sludge has high contents of calcium fluoride and water but low contents of iron, chromium and nickel. However, it would also pose a potential environmental problem if deposited outdoors over a long term because it contains heavy metals such as Cr and Ni. At present, solidification/stabilization (S/"

Jan 1, 2014

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

The paper discusses the special mixing requirements for Metallurgical Plant Design and outlines the problems associated with the agitation in Mineral Processing Industry today. It also covers new applications such as state of the art large leach tanks and backfill for deep mining.

Jan 1, 2004

By Raja V. Ramani, Ralph W. Barbaro

This paper presents an application of a Mixed Integer Programming (MIP) model to the problem of locating a centralized coal preparation plant and determining the production schedule for the mines. The system used for illustrating model application is hypothetical but realistic. It consists of three mines, two potential preparation plant sites, heavy media processing plant, and one market. The goal is to maximize the total profit of the system while satisfying the coal delivery schedule over a three year period meeting both quantity and quality requirements (ash, sulfur, and heat content). In addition to the base case, four different scenarios are analyzed. These four modifications are: (1) removing the flotation circuit from the preparation plant and sending all of the fines directly to the clean coal, (2) lowering the quality of coal from one of the mines, (3) applying discounts to the unselected site to improve its competitive position, and (4) adding a new mine. A detailed discussion of the base case and the effect of these four alternatives is presented.

Jan 1, 1983

One of the very simplest losses to overcome, and at small expense, is belt slip. To show the extent of the loss of money through slipping of the main belt alone, a chart, supplied us through the courtesy of the Cling-Surface Co., Buffalo, N. Y., is shown herewith, upon which this is easily ascertained. By glancing up and down column "D" of this chart it is evident that the "Cost of Avoidable Belt Slip per Year in Dollars" may vary all the way from the smallest sums into the thousands of dollars. Roughly, 2 per cent. of the potential power in the driving pulley of the engine is represented by "belt creep." Its loss is unavoidable because of the elasticity of the belt. This has been recognized and provided for in the chart, so the reader can go ahead and connect known values, while the intersection with column "D" gives the avoidable money loss. Simply zig-zag across the chart twice with a straight edge, as shown by the dotted lines and the money loss is immediately found. Thus if the total belt slip is 6 per cent. (column A) and the cost of coal per ton is $4 (column B), run a straight line through those two points and locate the intersection with column "C." Then from that point of intersection run over to the tons of coal burned per year (which in this instance has been stated as 1000 tons) and the cost of avoidable belt slip per year is shown by column "D" as being $160. To give a clear idea as to the meaning of "Total Belt Slip Per Cent." let us take an example. You find by means of a revolution counter that a given driven pulley is rotating only 940 times per minute. You figure that without slip it should rotate 1000 times per minute; 60 revolutions per minute, therefore, are absolutely lost. Dividing this 60 by the r.p.m. that pulley should make you get 0.060 or 6 per cent. which is the "Total Belt Slip." After having made this determination, it is a simple matter to apply it to the chart as has been done and determine the money loss per year due to such slip.

Jan 4, 1918

By Whitmell J. Smithwick, Charles E. Dandois

The first operating unit of the ring or annular shaft kiln as developed by Mr. Karl Beckenbach of West Germany was rated 185 tpd and was installed in 1962. Today, there are approximately 84 units in operation in Western Europe and Japan. Twenty-one additional units are under construction or on order. Sizes range from approximately 100-330 tpd, however, units of over 500 tons are in the design stage. The ring shaft kiln is a single vertical cylinder to which are attached the required service platforms, combustion chambers, injectors and distribution piping (Fig. 1). Auxiliaries such as recuperator, exhaust fan and dust collector are mounted on the appropriate service platform. The charging pan, feed conveyor, etc., are mounted at the kiln top and at the bottom of the kiln is the silo with a vibrator type discharger. Mounted separately is the required blower for injector air and fan for internal cooling.

Jan 4, 1973

By John Chipman, Peter J. Koros

The distribution of copper between liquid silver and liquid iron, Fe-C, and Fe-C-Si alloys was studied at 1600°C. From the data and the activity of copper in silver obtained from the phase diagrams, the activity coefficients of copper in iron up to 3 atomic pct are log y, = 0.90 — 2.4 N, and y": = 8.0. The effect of carbon up to saturation is expressed as log y = 1.8 Ni. The effect of silicon up to 11 atomic pct is insignificant. SLOWLY rising copper content of steel being produced is focusing attention on the lack of information on the behavior of copper in liquid steel. Chou' measured the activity coefficient of copper in liquid iron through its distribution between liquid silver and iron. He obtained a value of 9.12 at infinite dilution. Chipman' calculated the activity coefficient of copper in liquid iron at 1600°C from the phase diagram and obtained a value of 12. Ameiln and Pfeiffer- investigated the possibility of removing copper from liquid steel through the use of lead, silver, and bismuth as solvents. They also calculated that the activity coefficient of copper in liquid iron would be increased five or sixfold by addition of 4 pct C. Langenberg and Lindsay' are investigating the effectiveness of lead and sodium sulfide in removing copper from liquid iron. The experimental method was similar to that used by Chou' and Chipman and Floridis." Silver, copper, and iron were melted together, the copper distributing itself between the two layers so that at equilibrium a,.,, (in Ag) - a,.,, (inFe) and N,.,, (in Ag) yc,, (inFe) (in Ag). [21 N,.,, (inFe) Thus, from the concentration of copper in each layer and the activity coefficient of copper in liquid silver, the activity coefficient of copper in liquid iron can be calculated. The effect of an alloying element such as carbon on the activity coefficient of copper can also be determined from its effect on the distribution, provided the new element is essentially insoluble in silver. The one serious limitation of this method is the increasing mutual solubility of silver and iron in the presence of a third component. The mutual solubilities of pure iron and silver are negligible, but increase with increasing copper concentration beyond the range investigated. Ag-Cu System—The activity coefficient of copper in silver at 1600°C had to be estimated, since direct experimental data are not available. Kawakami" measured the heat of mixing of Ag-Cu alloys at 1200°C by a calorimetric technique. Scheil' calculated the heat of mixing in this system by use of the phase diagram, assuming that the solution is random and that the heat of mixing is independent of temperature. The terminal solid solution may be assumed to obey Raoult's law within the requirements

Jan 1, 1957

By J. A. Herdlick

"INTRODUCTION The BeVan quartz crystal prospect in the Indian Creek mining district of Lemhi County, west-central Idaho, was examined by an engineer of the Bureau of Mines 3/ on July 9, 1943. As quartz crystals possessing peizoelectric properties were on the critical list of strategic war minerals, further exploration of the deposit was recommended. The Bureau of Mines exploration began in December 1943 and continued through January 1944. Additional trenching was completed during May and June 1944. This report presents the data obtained by this exploration.ACKNOWLEDGMENTSIn its program of exploration of mineral deposits, the Bureau of Mines has as its primary objective the more effective utilization of our mineral resources to the end that they make the greatest possible contribution to national security and economy. The Mining Division, Lowell B. Moon, chief, conducts preliminary examinations, does the actual exploratoiy work, and prepares the final report. The Metallurgical Division, O. C. Ralston, chief, analyses samples and makes beneficiation tests.The quartz crystals found during the exploration were examined and classified at the Bureau of Mines College Park, Md., laboratory and checked at the National Bureau of Standards, Washington, D. C.Fred BeVan, of North Fork, Idaho, and Edward F. Cenis, of the Montana Coal & Iron Company, Spoane, Wash., provided much information on the history and past production of the property."

Mar 1, 1948

By C. Hosten, K. V. S. Sastry, D. J. Kawulok-Englund

Introduction Since the last decade the iron ore processing industry has undergone drastic changes from supplying screened and natural ores to manufacturing tailored concentrates and pellets. Increased need for the beneficiation of lower grade iron ore deposits and the increased demand for higher grade concentrates necessitate finer grinding of iron ores before concentration. The normal filtration stations used in the past are experiencing difficulties in dewatering the finely ground concentrates to moisture levels acceptable for balling. In addition to finer grinding, applications of complex beneficiation techniques has decreased the filterability of the concentrates and create problems in filtration of iron ore concentrates. An overview of the filtration literature indicates there are few research reports on the dewatering of iron ore concentrate slurries. There exists a need to understand the principles and mechanisms involved in the filtration of iron ore concentrate slurries with a goal of developing more efficient filtration techniques. This paper presents the results from the first phase of our research program that included the development of a reproducible benchscale filtration technique and the investigation of the filtration behavior of eight iron ore concentrates.

Jan 10, 1983

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Jan 1, 1938

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Jan 1, 1938

By E Amini, M Scott

Grade Engineering® involves the planning, integration and operation of coarse preconcentration techniques to reject uneconomic material as early as possible in the mining value chain to improve the quality of ore delivered to mineral processing activities. These preconcentration techniques include screening for natural and induced deportment of grade to finer size fractions during blasting and coarse breakage, sensor-based measurements of grade and mechanised sorting at particle or bulk scales and coarse gravity separation. Over the last decade, the Cooperative Research Centre for Optimising Resource Extraction (CRC ORE), with the support of its participants, has developed a framework for the holistic evaluation of economic and production impacts attributable to these preconcentration techniques. The framework combines the characterisation of preconcentration responses, geostatistical modelling, process modelling and simulation of downstream and upstream production impacts and the optimisation of strategic mine plans to quantify changes in net present value due to different preconcentration strategies. This framework has been applied to evaluate preconcentration techniques at numerous metalliferous operations and can also be used to quantify changes in the environmental metrics for energy, water and steel consumption and emission, tailings and waste rock production. This paper collates the results of these Grade Engineering® studies to show the range and magnitude of economic and environmental impacts attributable to preconcentration. The studies have shown that rejection of uneconomic material prior to mineral processing activities can have a pronounced improvement in energy, water and comminution steel productivity, emission intensity and the profitability of metalliferous mining.

Nov 10, 2020

By James H. Nulty, Joseph S. Roti

INTRODUCTION During a recent panel discussion on the topic of millwater antiscalants, the question was posed: "What would you like to see in the future?" Answer: 'We need an antiscalant that: can be fed without a pump (no maintenance), functions on residual basis (no fault), can be tested for colorimetrically in sub-part-per- mi 11 ion concentrations (no guessing) and will recycle back to the mill from tailings (no cost).” Impossible? Based upon today's level of technology, such a product does not exist. However, a few of the qualifications listed above are certainly food for future thought. Meanwhile, progress is being made in upgrading three essential elements in this critical area of ore processing. They are: 1. Antiscalant reagent research and development. 2. On-line monitoring programs. 3. Reagent storage and feed systems. The objective of this paper is to out- line the current direction being taken in each of these areas and to present case histories as examples of how theory has been applied in field situations to better improve operations. LABORATORY SCREENING TESTS Recent technical papers have explained the various generic chemical raw materials which are available in the marketplace. While the merits of each chemical component are open to debate, the processes used by most major reagent suppliers to evaluate products in the laboratory are fairly standard. Common techniques incorporate both static and dynamic conditions in the search for the most cost effective product. Static Test Perhaps the most common test procedure used is based upon NACE Standard Test Method 03-74, which is "a laboratory screening test to determine the ability of scale inhibitors to prevent precipitation of calcium sulfate and calcium carbonate from solution." The objective of this test is to measure the change in calcium ion concentration in a container of saturated stock solution prepared from calcium chloride and sodium carbonate which is heated in a water bath. After 12 hours at constant temperature (160°), the mother liquor is pipetted into a beaker and a titrated for gm/l of unprecipitated calcium ion. A simultaneous test is performed on stock solution with antiscalant added. The results are then charted to determine the most effective antiscalant (with respect to preventing precipitation of calcite). It should be noted that this procedure examines only the ability of a chemical additive to retain calcite in a saturated solution where thermal stress is applied. This is commonly referred to as sequestration or stabilization. It has been documented that products which act in this manner (i.e., inorganic and organic phosphates) are good scale inhibitors. However, low molecular weight anionic polymers function in a much different manner by distorting calcite crystal size, shape and charge density, thus preventing scale formation. Because of this different mode of operation, static tests may not prove conclusive where polymers and sequestrants are being compared. The simplicity of this test makes it easily adaptable for field use, although modifications are often made.

Jan 1, 1986