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By F. F. Peng

The raw-coal feeding system and the upward medium current flow in a dense-medium vessel (DMV) are, in addition to the residence time of particles in the separator, important factors for efficient coarse particle separation. The feed material from a raw coal screen and/or pre-wet screen merge with the major volume of circulating medium as push medium, guided by an adjustable submergence baffle plate deep fed into the separator for effective separation. The remaining volume of the circulating medium enters from the purge and drain hoppers at the bottom of the DMV. This generates a gentle upward medium current flow in the separator, which prevents dense medium stratification/settling and merges as part of the push medium. A two-dimensional, two-phase model developed using computational fluid dynamics (CFD) is used to describe the flow pattern and evaluate the function of the dip plate for understanding the mechanisms of particle separation in the separator. The particle movement in the DMV is also tracked to investigate the major operating parameters affecting the separation performance of the DMV. The simulation results of the DMV separation performance are validated by in-plant test data.

Jan 1, 2007

By T C. Ooi, L Lu

Owing to ongoing development of new orebodies and changes to existing mining operations, new ore types with distinct mineralogy, porosity and textures are generated as feed for iron and steelmaking. Moreover, further product differentiation occurs as fine ores are increasingly beneficiated to achieve grade. A study was necessary to characterise a suite of iron ore fines in terms of their propensity for formation of quasi-particles and green bed permeability prior to sintering. In this study, two Marra Mamba bedded iron deposit (BID) samples, one Pisolitic channel iron deposit (CID) samples, two dense hematite samples (coarse and fine grained microplaty hematite), one hematite-goethite BID samples and one hematite-goethite bedded/detrital iron deposit (BID-DID) composite sample were investigated. Both Marra Mamba fines and Pisolitic fines exhibit the highest green bed permeability followed closely by hematite-goethite fines, fine microplaty hematite fines, hematite-goethite BID-DID composite fines and coarse microplaty hematite fines at a moderate sintering bed depth of 500 mm. While porous fines generally form better quasi-particles upon granulation than dense fines, the amount of moisture addition increased due to the moisture saturation of the porous fines. As sinter strands become ever larger and deeper, it was necessary to examine the granulation performance of these ore types at a range of bed depths ranging from 300 to 900 mm. On average, the green bed permeability achieved by different ore types, decreased by approximately 7.5 - 9 Japanese Permeability Unit (JPU) as the bed height was increased from 500 mm to 900 mm. In order mitigate the effect of lower bed permeability of the deeper beds, small additions of hydrated lime were investigated as a binding agent to facilitate the formation of strong micro-pellets. With the addition of three per cent hydrated lime, equivalent to approximately 2.3 per cent burnt lime, the majority of fines achieved adequate permeability for deeper bed sintering, with the exception of coarse microplaty hematite fines. In simulated coastal steelmaking sinter blends, it was found that the overall granulation performance of the blend is contributed to by individual fines constituting the blend. Therefore fines possessing favourable granulating properties impart better granulation to the blend as a whole.CITATION:Ooi, T C and Lu, L, 2013. Effects of ore type and binder addition on the formation of quasi-particles and green bed permeability at different sintering bed depths, in Proceedings Iron Ore 2013 , pp 459-466 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

By David R. Peterson

This user's manual covers a set of 5 computer codes which were generated during development of a simulation of deep ocean mining operational economics. Table 1.1 provides a summary of these codes, and a more thorough description of each follows, In addition to the descriptions in this user's manual, internal documentation is provided throughout each code, except for the very simple ones, for ease of understanding or possible modification. The two major codes are MINER, which performs the deep ocean mining system operational simulation, and MSWEEP, which calculates the sweep and feed efficiencies for a set of collector paths generated by MINER. The standard operating procedure is to input the system and operating parameters to MINER, which computes a set of collector paths automatically and stores the information in the fife PATH.DAT. MSWEEP subsequently retrieves the collector path data in PATH.DAT, and uses it to calculate sweep and feed efficiency. Separating MINER and MSWEEP allows parmetric execution of MSWEEP without recomputing the collector paths with MINER. PIPEST, RESP, and OBSTCL are special-purpose codes described in more detail in the following documentation.

Jan 1, 1979

SME's Annual Membership Report - 1961 Membership Directory

Jan 7, 1961

By W. G. Pfann

A simple method of obtaining multistage batch separations by crystallization was described recently. Known as zone-refining, it comprises passing short molten zones through a long solid charge. This technique can now be used on a continuous basis by means of the zone-void method described in this paper. Feed enters, at an intermediate point, a column down which molten zones travel, and waste and product leave at the ends. Materials move in the column through the agency of voids, which are introduced at the ends and travel toward the feed inlet. The voids and molten zones are moved by external heaters in a simple manner, and the principle of reflux is utilized. ANEW method of obtaining multistage separations by crystallization was described recently.' Named zone-refinina.,-, the method comwrises slowly passing a series of molten zones through a long solid charge. Solute becomes concentrated at one or the other end of the charge, depending on whether it raises or lowers the freezing point of the solvent. The separation increases with the number, P, of zone-passes, approaching a limit as P approaches infinity. Zone-refining has been highly effective in purifying germanium and other substances.2,3 and new applications are steadily increasing. Zone-refining is a batch method and as such it has certain limitations inherent in batch operation. If it could be made continuous, its scope and utility would be greatly broadened. This end has been achieved by the zone-void method described in this paper. In the zone-void method feed is introduced continuously at an intermediate point in a column down which molten zones travel, while impure waste and purified product leave at the ends. Both the flows of feed, waste and product, and also the travels of the zones, are actuated by external moving heaters in a simple manner; and the system utilizes the principle of reflux. The method provides, in the field of crystallization, the counterpart of the continuous fractiona-tion column in the field of distillation. The following will be discussed: apparatus and mode of operation, fundamental nature of the separation, design theory, and practical considerations. The method will be described in terms of a binary solute-solvent system in which the solute is an impurity to be removed and the solvent is the desired product. The distribution coefficient, k, defined as the ratio of solute concentration in the solid freezing out of a molten zone to that in the liquid in the zone, is assumed to be constant and less than one. The process is equally effective for k's greater than one and for ternary or higher order systems. Method and Apparatus The essential features of a continuous zone-refining process are represented in highly generalized form in Fig. 1. A series of molten zones, produced by moving heaters, travels slowly down the column or charge (to the left in Fig. 1). If there were no flows of feed, waste, or product, the process would simply be batch zone-refining, the action of the molten zones being to sweep solute down the column, solvent up the column. For the process to be con- tinuous, with stripping and enriching sections in the column, feed must enter, and waste and product must leave, as indicated. The zone-void process accomplishes both objectives, namely, the indicated movements of zones and the indicated flows of material. Zones are moved by moving heaters, just as in batch zone-refining. Materials are made to flow by creating voids at the waste and product exits and causing these voids to move to the feed inlet. Since there must be a net flow of material from the feed inlet to each of the outlets, the indicated movements of voids are in the desired directions, because movement of a void in a given direction corresponds to flow of material in an opposite direction. In order to produce the desired movements of voids, the column is folded into two vertical sections having the feed inlet in common at their upper ends. Voids are displaced upward by the liquids in the molten zones and their travel is actuated by the motions of the zones. Voids travel with the zones in the enriching section and move continuously. Voids travel opposite to the zones in the stripping section and move intermittently. Creation and travel of voids will now be examined in detail. The enriching section of a column in operation, with its void generator, is shown in Fig. 2. The column section is a vertical tube around which a series of closely fitting, regularly spaced heaters travel slowly upward. Each heater produces a molten region, the temperatures of the heaters and the cooling between heaters being controlled so as to maintain the molten zones approximately constant in size. A void is normally present atop the molten zone in each heater. As the heater rises, it continuously melts solid above it, which drips through the void into the molten zone and continuously freezes out solid below it, of concentration k times that of the liquid in the zone. When a molten zone and void reach the feed inlet, which is kept molten, the void is displaced by an equal volume of feed liquid. Generation of voids of controlled size in the enriching section is shown in Fig. 3. The void generator is a tube of small cross-section, provided with lateral heat-conducting fins which sense the position of the heater. Liquid can escape only when the entire outlet tube is within the heater. If any part of the outlet tube is outside the heater, liquid cannot escape,

Jan 1, 1956

By Walter E. Lewis

The Cuyuna iron range in Crow Wing County, Minn., is the principal producer of manganiferous iron ore in the United States. The ore is mined from a manganiferous iron formation in the bedrock of the Cuyuna district. This iron formation is divided into two general lithologic types-thin-bedded and thick-bedded. The unoxidized, thin-bedded iron formation is composed of iron silicate with associated fine-grained carbonates of iron and manganese. The thick-bedded iron formation carries iron and manganese oxides with intimately associated silica. Oxidation and leaching of the original iron formation have been the most influential factors in producing manganiferous iron-ore concentrations in quantities large enough to exploit on a commercial basis. The unoxidized and the partly oxidized and enriched iron-formation materials are suitable as a source of ferrograde manganese ore. These materials have been classified according to their amenability to mineral dressing and degree of oxidation and leaching.

Jan 1, 1958

By Henry M. Howe

PART I. § 1. Introduction. After giving certain definitions, this paper gives the reasons which led to Roozeboom's form of the diagram of the freezing-point curves and general equilibrium of the iron-carbon compounds, steel and cast-iron ; and then those which led to replacing it with the present or double diagram, in which graphite is at all temperatures held to be the more stable and cementite the less stable (metastable) form of carbon. Next in § 10 the evidence supporting this greater stability of graphite is examined, then in § 17 that which seems to oppose it, and in § 22 a summary of the evidence is presented. The second part of the paper considers the topography of the graphite-iron diagram, rejecting summarily the solubility-lines based on the data of Charpy and Grenet and of Mannes-mann, on the ground that the data even on their face do not really point toward these lines, and that they are wholly incompetent to determine any solubility-lines whatsoever, because they report together with the dissolved carbon also an indeterminate quantity of cement-carbon, and because even such carbon as is actually dissolved, only in part represents the solubility of graphite, and in another and indeterminate part represents the solubility of cementite, or at least of carbon in presence of cementite as distinguished from graphite. It might be thought that graphite is but slightly soluble in iron, because of the absence of any suggestions of eutectoid graphite, corresponding to the eutectoid cementite of pearlite in very many cases in which the cast-iron seems to consist of ferrite and massive graphite only, and because of the micrographs of Goerens and Gutowsky. In these the primary austenite of cast-iron quenched but little below the eutectic freezing-

Jan 7, 1908

By Gary B. Petersen

The author has developed a seven-step process called the Applied Rock Mechanics (ARM) Protocol that the author has used to solve numerous roof stability problems in mines throughout North America since 1975. It is a practical evaluation process using trained observation, innovative data collection instrumentation, video borehole cameras, analytical processes, and controlled testing methods that have become highly effective in solving ground control problems in every type of underground mine. This paper will describe each of the steps of the protocol and how the rock mechanics engineer applies them to resolve a typical ground control problem.

Jan 1, 2012

By A. L. Peyton

Manganese-oxide minerals occur in irregular pockets distributed through the residual clays of the Cartersville district, which lies 45 miles north-west of Atlanta in Bartow County, Ga. The district produces, or has produced, limonite, barite, specular hematite, ochre, ferruginous manganese, manganiferous iron are, as well as manganese. Bureau of Mines engineers have visited the district on three occasions - in December 1939, February 1941, and October 1942. During the visit of February 1941, 80 tons of samples of manganese-bearing clays were taken from seven properties for ore-dressing tests, which were conducted at the Bureau of Mines laboratory at Rolla, Mo. The. October 1942 visit was for the purpose of reappraising the district in the light of knowledge gained from the ore-dressing studies and to prepare a plan for the district's development. Exploratory drilling was begun in March l943 and was continued until June 1945. Nine properties were explored that, roughly follow a line beginning near Cartersville and extend to a point 2 miles northeast of the village of White. The work consisted of bulldozer trenching, rotary bucket churn drilling, sampling, and analytical determinations. The manganese district is 18 miles long with a maximum width of 7 miles at Cartersville, which lies at the center of the district. The work of the Bureau was confined to the northern half of the district between Cartersville and Pine Log. The properties explored comprise but a small portion of the Potential producing area.

Jan 1, 1949

By William H. Waggaman

The United States not only is the largest producer of phosphate rock but also manufactures and consumes greater quantities of phosphate products than any other country. New and expanding uses will increase even further the demand for this basic raw material. Only a small proportion of the total quantity of phosphate rock mined and marketed is consumed directly in a raw or unprocessed state; the vast bulk is chemically treated and converted into [ertilizer] and a wide variety of finished products that play a vital role in [r] national economy. Three fundamental methods are employed in decomposing phosphate rock. 1. Reduction by carbon at high temperatures. 2. Treatment with one or more acids. 3. Calcination or fusion with silica, sodium salts, magnesium minerals, or small quantities of phosphoric acid.

Jan 1, 1960

By L. D. Jaffe, R. K. Pitler

IT would clearly be advantageous to produce molten titanium, suitable for alloying and casting, directly from the relatively inexpensive tetra-chloride, without using a metallic reducing agent. Accordingly, a preliminary investigation has been made of the production of titanium by hydrogen reduction of the tetrachloride in an electric arc furnace. For the reaction: TiCl4(g) + 2H2(g) ? Ti(s) + 4HCl(g), Lockhart and his colleagues' have calculated equilibrium constants at 800" to 1500°K, based on tables of Brewer.' Extrapolating these data on the assumption that the heat of reaction varies linearly with temperature, and using Brewer's value for the heat of fusion, there is obtained for the reaction: TiCl4(g) + 2H2(g) ? Ti(l) + 4HCl(g) an equilibrium constant at 2100°K (100" above the melting point) Kp = (PHCl)4/pTlcl4 (Ph2)2 = 0.167 atm. If this is correct, starting with hydrogen saturated with TiC14 at 25°C, where the vapor pressure of the latter3 is 12.6 mm, and maintaining the overall pressure at 1 atm, 99 1/4 pct of the TiC14 would be reduced to Ti at equilibrium. An arc furnace designed for the melting of titanium and lined with water-cooled copper was used for the preliminary experimental check. Its inside diameter was 2 7/16 in. and height 6 in. The water-cooled electrode was tipped with 3/4 in. diam tung-sten. In the one run carried out, the furnace was charged with a starting batch of 293.35 g of scrap titanium,' analyzing 99.24 pct Ti. After the system was twice pumped out and flushed with argon to remove air, the arc was struck in argon and kept at 400 amp dc, electrode negative, for 1 min to melt the starting batch. Commercial tank hydrogen, dried by passing over silica gel and bubbled through technical grade TiC14 at room temperature, was then admitted to the furnace. The inlet tube directed this gas downward toward the arc and melt; the gas outlet was at the top of the furnace. The quantity of gas passed through was not measured accurately but is estimated to be about 50 liters per min at a few mm of mercury above atmospheric pressure. The arc was held in the H2-TiC14 mixture at 400 amp with little difficulty. After 10 min the arc burned through the copper lining above the melt, admitting water and ending the run. (This had happened frequently in straight melting runs under argon, and is not thought due to the atmosphere used.) In the furnace were found an ingot, several small pieces, and a little metallic powder, evidently formed when water entered the hot furnace. Some titanium also clung to the electrode. The portions were weighed and separately analyzed for titanium. Their total weight was 321.76 g of 94.82 pct overall purity. Much of the impurity consisted of tungsten broken off when the titanium was removed from the electrode. No detailed impurity analysis was made because the starting materials were impure and flooding the furnace introduced further contamination. Thus, the metal recovered contained 305.08 g Ti, as compared to 291.12 g in the starting batch. Titanium could not have been carried over from previous melting runs in the furnace, as the interior was sandblasted and the electrode ground clean before the run. Evidently at least 13.96 g of Ti were produced from TiCl4. Since this would require complete reduction of 430 liters of TiCl, gas at 12.6 mm of Hg, the efficiency of the process must have been rather high. Experimental work is continuing. No reason is seen why the process could not be scaled up to the size of a 30-ton steel-melting arc furnace. In a large unit there would be a large molten pool from which liquid metal could be withdrawn for alloying and casting. The problems of electrodes and power supply are similar to those arising in arc furnaces for melting titanium. Unreacted TiC14 and the HC1 in the exhaust gases could be recovered; unreacted H2 could be dried and re-used. The melt would be saturated with hydrogen, whose solubility in liquid titanium is unknown. Small amounts of hydrogen seem to have little effect on the properties of titanium.5 If the amount introduced was so large as to impair properties or cause casting difficulties, it could be reduced by holding the melt under an inert gas or by pumping on it, prior to casting. Acknowledgment The authors would like to thank I. Preble and M. Goldman for their valuable aid in the experimental work and the thermodynamic calculations, respectively. References 1R. J. Lockhart, J. J. Ward, M. J. Hussey, and J. W. Clegg: In unpublished report, "U. S. Air Force Project Rand. Titanium and Titanium Base Alloys." Battelle Memorial Institute (1949) 37-44. 2 L. Brewer et al: Chemistry and Metallurgy of Miscellaneous Materials—Thermodynamics. Ed. by L. L. Quill. pp 13-39, 60-192. (1950). McGraw-Hill Book Co. 3 K. Arii: Scientific Reports. Tohoku Imp. Univ. (1933) 22, 182. 4 R. S. Dean, J. R. Long, F. S. Wartman, and E. L. Anderson: Transactions AIME (1946) 166, 369-381; Metals Technology (Feb. 1946) TP 1961E. * R. I. Jaffee and I. E. Campbell: Transactions AIME (1949) 185, 646-654; Journal of Metals (Sept. 1949) TP 26813.

Jan 1, 1951

By David Rahal, Des Erasmus, Ken Major

"Over the past twenty years there has been an increased emphasis on fine grinding in the minerals industry. This spurred the development of a range of new fine grinding equipment in the 1990s. At that time, the Knelson-Deswik mill was developed in South Africa for the pigments industry.Knelson Milling Solutions currently offers a wide range of laboratory, pilot and production scale mills. These mills bridge the gap between traditional low speed vertical mills and the high speed horizontal mills (tip speed < 3 and > 15 m/s respectively). They are designed to run at tip speeds between 10 and 12 m/s and are capable of operating with very dense grinding media. The result is a power intensity that is higher than the low speed mills and overlaps that of the high speed mills. This allows the mill to be customized for a wide range of grinding applications.A case study at Aquarius Platinum’s Kroondal Mine shows that the Knelson-Deswik technology can produce a consistent product size despite feed variations. It also shows that the throughput was proportional to the slurry density at Kroondal. The case study identified increasing the slurry density as a process change that can increase throughput and reduce the specific energy consumption.INTRODUCTIONThe increasing complexity and finer grain sizes of modern ore bodies have increased the importance of effective mineral liberation in obtaining acceptable grades and recoveries (Jankovic, Valery and La Rosa, 2003; Lichter and Davey, 2006; Napier-Munn, Morell, Morrison and Kojovic, 1996). This trend has increased the significance of fine grinding technology because the input energy increases exponentially as the required product size decreases from 100 to 10 microns (Gao and Weller, 1993a).The US Department of Energy recently provided the context for the potential energy savings in the US Mining Industry (BCS, 2007). They reported that grinding consumes approximately 494 TBtu/year. This was over double the consumption of the second highest equipment category, diesel materials handling systems (211 TBtu/yr). The proportion of grinding energy allocated to fine grinding can only be expected to increase as finer ore deposits are mined. It is therefore critical that fine grinding efficiency continue to improve through the selection of appropriate equipment and the development of best practices for existing applications."

Jan 1, 2011

By Ted J. Williams

Underhand cut and fill mining methods are used at the Lucky Friday, Stillwater, and Galena mines in the western US in an effort to safely mine in difficult ground conditions and as a primary mining method where access below the ore is limited. The method has evolved at each mine to meet the unique situations associated with their milling procedure and mine layout. Details of the operational aspects of the backfilling and mining cycles at each mine are presented in this paper, accompanied by a summary of the mechanical properties of the backfill and its in-situ geomechanical behavior.

May 1, 2007

By Henry Kreisinger, WALTER T. RAY

General statement.-The briquetting tests conducted by the technologic branch of the United States Geological Survey had their beginning in the testing of coals and lignites at the Louisiana Purchase Exposition, St. Louis, Mo., in 1904. Briquetting tests were made at the fuel-testing plants at St. Louis, Mo., and at Norfolk, Va. Similar tests to be made at the experiment station of the Bureau of Mines at Pittsburg, Pa., form an essential part of investigations looking to the determination of the fuel value of the coals and lignites belonging to or for the use of the United States. The tests have comprised (1) the manufacture of briquets to determine the adaptability of different coals and lignites to the process and the merits of different binding materials; (2) physical tests to establish the fitness of the briquets to withstand weathering, transportation, and handling; and (3) steaming tests to prove the calorific value of the briquets in boilers of different types used by the Government, and, by comparison with the raw coals, the benefits to be derived from briquetting. Details of the tests have been issued in other publications. list of Government publications on fuel testing and briquetting is given at the close of this volume. Tests at Norfolk.-The fuel tests conducted at Norfolk included a detailed investigation of a number of Virginia and West Virginia coals that are bought by the United States Government for the Navy and for use in constructing the Panama Canal, and are extensively used by the merchant marine, manufacturing plants, and railroads. Through a cooperative arrangement with the Navy Department steaming tests of the coals were undertaken to determine the relative merits of the same coal when burned raw or as briquets in marine boilers. Preliminary arrangements were made between Joseph A. Holmes, then expert in charge of the technologic branch of the United States Geological Survey and now director of the Bureau of Mines, and Rear-Admiral Charles W. Rae, Chief Engineer of the United States Navy. The tests were made on board the U. S. torpedo boat Biddle, designated for the purpose, beginning December 6, 1907, and ending January 27, 1908. The coal used in these tests came from the Sewell and Beckley beds in the New River district of West Virginia. With the particular equipment used in the tests both coal and briquets were far. from smokeless; consequently the data of this bulletin are applicable only by analogy to parallel operation with a coal more nearly smokeless, but nevertheless applicable with much reliability. The possibilities of coals of different composition are indicated by the data to be published in a bulletin of the Bureau of Mines, wherein the present authors describe a number of tests in which only a very slight amount of smoke was emitted while burning raw coal from the Pocahontas No.3 bed, West Virginia, and briquets made therefrom at rates of combustion very much higher than any rates used in the torpedo-boat boiler. At a combustion rate of 120 pounds of the Pocahontas briquets per square foot of grate surface there was scarcely any smoke. It was the original intention to make a set of preliminary steaming tests alongside a dock (which tests furnish material for this bulletin) and to finish with a set of running tests at sea; but the running tests were never made, for lack of time and men.

Jan 1, 1911

The Annual General Meeting of the South African Institute of Mining and Metallurgy was held in Kelvin House on Wednesday, 25th August, 1976. Dr R. E. Robinson, (President) was in the chair, and declared the meeting open at 16h00. Obituaries The president: It is my sad duty to announce the death of the following members of the Institute : P. A. Creewell, Life Fellow; A. S. Davis, J. S. Hay, P. O. Large, N. McCaskell, and H. G. M. McKerrow, Life Members; C. L. Bererge, Fellow; T. Bulloch, G. Douglas, R. W. Forrester, and C. G. P. Sack, Members; P. R. Garcia, W. A. Naude, and J. P. Slager, Associates. As a mark of respect to the memory of the deceased and in sympathy with the bereaved, I ask you to rise and observe a few moments&apos; silence. Minutes The President: May we confirm the minutes of the General Meeting held on March 17th, 1976, as published in the June issue of the Journal? Agreed. Welcome The President: It is my pleasant duty to welcome a number of guests and distinguished visitors this evening. First, I would like to welcome Mr R. Plumbridge, the President of the Chamber of Mines, and Mrs Plumbridge. We have had very close co-operation with the Chamber of Mines, and it is particularly gratifying that Mr Plumbridge has had time to be with us this evening in spite of very heavy engagements. Ook &apos;n spesiale woord van verwelkomirg aar mrr. N. Orsmord, die Staatsmyringerieur, en mev. Orsmord. Ors wil hierdie geleertheid gebruik om mrr. Orsmord geluk to wens met sy aarstellirg as die Staatsmyrirgerieur en die hoop uitspreek dat dit baie jare van samewerkirg met hierdie Irstituut sal beteker. We also welcome guests who, in one way or another, have been able to attend most of our Annual General Meetings, the Honourable I. B. Dillon, the Minister of Mines of Rhodesia, and Mrs Dillon. Once again we welcome them most heartily to our Annual Meeting. I would also like to welcome the representatives of many of our sister institutions : the Honorary Treasurer of the AS & TS, Mr D. G. Maxwell; the President of the South African Institution of Mechanical Engineers, Mr A. R. Adams; the Vice-President of the South African Institute of Electrical Engineers, Mr K. A. H. Adams, and Mrs Adams; the President of the Institute of Certificated Mechanical and Electrical Engineer s, Mr Martin and Mrs Martin; a member of the Executive of the South African Institute of Civil Engineers, Professor G. Bright ; the President of the South African Chemical Institute, Mr Vic Shuller, and Mrs Shuller; the President of the Institute of Welding, Professor F. P. A. Robinson, and Mrs Robinson; the President of SACPE, Mr M. R. Gericke ; the President of the Institution of Metallurgists, South African Branch, Mr D. R. Allen; the President of FSPE, Dr P. J. D. Lloyd, and Mrs Lloyd ; Mr A. J. H. Pole of the Association of Mine Managers, and Mrs Pole; and Mr M. D. Hoskin of the South African Institute of Foundrymen. A special word of welcome to Mr G. C. Young of our Orange Free State Branch, and Mrs Young ; and to Mr E. J. Jackson of our Witbank/Middelburg Branch. We have great pleasure in welcoming Dr F. M. Naude, the Scientific Adviser to the Prime Minister, and Dr W. C. J. van Rensburg, Technical Director of the Minerals Bureau. The latter is a new organization, and we hope their attendance here this afternoon forecasts a long period of co-operation between the Minerals Bureau and this Institute. Ook wil ek hartlik hier verwelkom die Rektor van Die Randse Afrikaanse Universiteit, professor G. van N. Viljoen. Ors het baie noue samewerking met die Universiteit gekry, en ek wil ook bier verwelkom die twee verteenwoordigers van die Geologiese-afdeling en die Chemie-afdeling, naamlik, dr C. Roering en professor G. W. Holzapfel. Last, but by no means least, a word of welcome to a very old friend, Mr Eric Boden, the Manager of AS & TS. To all our guests, a very hearty welcome to this our Annual General Meeting. Membership The President: It is required by our Constitution that I announce the names of a number of candidates who have been elected to various grades of membership. The list is a very long one, and I suggest that we simply record that these names will be published in the Journal. Any one who wishes to comment on these members should do so in writing to our Secretary. Agreed. Fellows: G. Paterson, P. H. Adler, F. W. G. Schoning, J. P. de Villiers, F. I. de Waal, J. R. H. Shaw. Members: C. J. Slavin, P. J. Perkins, R. J. I. Butler, A. M. Hattingh, C. E. Alvey, J. A. Engelbrecht, P. L. Gouws, J. P. Martin, J. F. Taylor, K. A. G. Blendulf, A. R. Gibb, W. E. Smith. Associate Members: D. R. Broderick, D. F. Cilliers, D. N. Stuart, P. Kraus, B. G. Pinder. Associates: D. E. Gibbon, D. F. Jacobs, R. F. W. Swanepoel, R. D. H. Bebb, H. H. Vernall, R. D. Gibson, J. P. van Wyk. Graduates : A. N. Morris, D. J. Payne, G. S. Worthington, A. S. Dennison, A. R. Godfrey, N. Holloway, W. J. Rankin, G. S. Lyndon, R. J. van Proctor, J. J. Versluis, G. M. Miller. Students: B. J. Rabie, G. D. Webber, W. L. Blair, I. D. Mathews, K. McVey, S. W. Beer, R. W. J. Bentley,

Jan 10, 1976

By Karl J. Benecke

Though the first patent on a bucket wheel excavator was granted in 1881 in the U.S. this technology was developed in Germany to the high standards of today. However, this development was only possible in parallel with the development of shiftable belt conveyors and boom stackers. The combination of these three types of equipment in the open pits of the Federal Republic of Germany has made lignite the cheapest energy source for power generating, despite the low heating value of German lignite in the order of 8500 kJ/kg (3600 BTU/lb) only. Comparing the costs for fuels used in power generating on the basis of their heating value, the relation between lignite, high quality black coal, and imported crude oil is in the order of 1 to 1.5 to 2.5, hence highly in favour of lignite. These cost advantages for lignite in Germany are due mainly to the concentration of lignite mining in a few high-output open pit mines. The largest one, Fortuna Mine in the Cologne district, produced a, total of 45 000 000 t lignite in 1976. That equals the annual fuel need of a 5750 MW power plant running 7000 hours per year at full capacity. Altogether 95 200 000 bank m3 equal to 200 000 000 t overburden had to be removed to accomplish this lignite output. The largest BWE (Fig. 1) ever built is working in the Fortuna Mine, too. At a digging height of 51 m above crawler level, the total vertical reach in relation to the face conveyor level is 98 m by means of its conveying bridge and loading unit. The hourly output is 10 400 bank m3. Its weight in operation is 13 000 t, its wheel diameter 21.6 m, as high as a 7-storey building.

Jan 1, 1977

By M R. Asef

Although generation of environmentally harmful hydrogen sulfide is considered to be related to sour gas fields, sweet gas reservoirs and waste fluids, travertine springs and geothermal fields may also produce acid gas (H2S and CO2) under certain conditions. The potential parameters influencing such a possibility need to be fully investigated to evaluate the extent of the consequences and predict hydrogen sulfide generation accordingly. Reports on hydrogen sulfide production at different regions and fields were gathered worldwide for the cases when some of them did not have such a problem previously. In order to inspect different aspects of these case studies, some other regions with no hydrogen sulfide content were selected as control. In all cases water samples were taken carefully for chemical analysis in the laboratory. Besides, mineral compositions were characterised in some cases for better evaluation. In order to give a thorough prospect of hydrogen sulfide production, a geochemical model was prepared. In situ engineering geological conditions, especially pressure and temperature, were assigned in the model to investigate the possibility of H2S generation using PHREEQC software. The results of modelling revealed that although many similarities existed between mineral compositions of the cases, certain formations were particularly prone to hydrogen sulfide generation. For better evaluation, important exchangeable cation ratios such as potassium/sodium, calcium/magnesium and ion proportions like sulfate and chloride were calculated as trends of the formations. As expected, the model could successfully monitor a higher ratio of SO42- and Cl- in potential H2S generator formations. Wide variation in percentage of HCO3- was considered as a secondary observatory parameter for better estimation. In addition, it was noticed that higher anion content created an environment ready to absorb any cations and this makes H2S generation easier. Finally, using correlations between predictors and estimators for low depth and deep samples defined the degree of effectiveness for every sample.

Sep 26, 2011

By Stephen R. Iverson, Jeffrey C. Johnson, Donovan J. Benton, Michael J. Raffaldi, Lewis A. Martin

"The NIOSH ground control safety research program at Spokane, Washington, is exploring applications of photogrammetry to rock mass and support monitoring. This paper describes two ways photogrammetric techniques are being used. First, photogrammetric data of laboratory testing is being used to correlate energy input and support deformation. This information can be used to infer remaining support toughness after ground deformation events. This technique is also demonstrated in a field application. Second, field photogrammetric data is compared to crackmeter data from a deep underground mine. Accuracies were found to average 8 mm, but have produced results within 0.2 mm of true displacement, as measured by crackmeters. Application of these techniques consists of monitoring overall fault activity by monitoring multiple points around the crackmeter. A case study is provided in which a crackmeter is clearly shown to have provided insufficient information regarding overall fault ground deformation. Photogrammetry is proving to be a useful ground monitoring tool due to its unobtrusiveness and ease of use.INTRODUCTIONNIOSH Mine Safety ResearchPhotogrammetry systems have been implemented by the National Institute for Occupational Safety and Health (NIOSH) as part of its ground control research to improve mine safety. Conventional monitoring of ground movement has typically focused on movement of a few discrete points that are marked or anchored to instruments. Loss of a designated point or anchor, or the ability to identify a stable point in an unstable area, have often frustrated monitoring efforts. NIOSH researchers are using photogrammetry to conduct full-field measurements of rock surfaces underground. Periodic measurements of the entire surface of a ramp allow patterns of displacement to be observed. Full-field measurements provide much more insight into ground behavior than point measurements.BackgroundPrevious work by Benton, et al. (2014) discussed laboratory calibration and verification testing of two photogrammetry systems being used by NIOSH researchers. Both systems utilize stereoscopic image pairing techniques for photogrammetric reconstruction. This paper discusses how a laboratory system was used to conduct volumetric analyses of testing of shotcrete panels with varying types of reinforcement. The laboratory system was found to be accurate to within 2 mm. Additionally, a field photogrammetry system was found to produce linear measurements within 1 mm of known lengths in laboratory conditions. This system is currently being used at a deep underground mine to observe ground support conditions and deformations that have occurred. This paper discusses comparison of preliminary volumetric measurements of rib deformation to laboratory tests, as well as comparison of the field system against a crackmeter installed across a fault surface where it intersects the ramp."

Jan 1, 2015

By K. H. J. Clarke

"ALTHOUGH nickel was not identified as a separate element until a Swedish scientist, Axel Frederick Cronstedt, did so in 17 51, its utilization dates from pre-historic times. From its humble beginnings nickel has grown in importance until, today, in our clocks, television sets, automobiles, and air-planes, in our bridges, tableware, farm implements, and the equipment which processes our food, it is constantly at our service.For convenience in assembly, data and information relative to The Marketing of Nickel will be outlined in five phases: (A) Sources, (B) Distribution, (C) Utilization, (D) Market Development, and (E) Conclusions."

Jan 1, 1960

By P. Latostenmaa, O. Forsén, M. Lundström, J. Aromaa

The history of electrolytic copper production in Finland is closely connected with the industrialization of Finland and events of the Second World War. Production of cathode copper by electrowinning from the processing of Outokumpu ore started in 1913 based on the Hybinette nickel process, however due to the poor performance of this process, the plant was closed in 1929. Outokumpu commenced copper smelting at Imatra in eastern Finland in 1936 producing blister copper for refining outside the country. Due to the threating military situation, a new electrolytic copper refinery was built in 1941 at Pori in western Finland to refine the anodes. This was the first electrolytic copper refinery in Finland. It was built with 108 cells having a capacity of 12,000 tonnes/y. In addition, the smelter was relocated to nearby Harjavalta in 1944–1945. To fulfill the war reparations included in the peace treaty with the Soviet Union, a doubling of the smelter capacity was realised, with the Pori refinery capacity increasing to 20,000 tonnes/y in 1946 by adding additional cells. The cornerstone of Finnish metallurgical technology — Outokumpu flash smelting — was invented with the new technology commissioned at Harjavalta in 1949 replacing the electric furnace and providing a further increase in copper output with subsequent enlargement of the refinery. By 1963, the plant consisted of 380 cells with a capacity of 48,000 tonnes/y of cathode copper. The Pori electrolytic copper refinery has operated continuously since 1941. Commencing in 2004 the Pori refinery has been part of the Boliden group of Sweden. The most recent plant upgrade was in 2007 with the installation of permanent cathode technology. In 2017, the production of the Pori refinery was 133,000 tonnes of copper cathode, 2,300 tonnes of nickel sulphate, along with 66 tonnes of silver and 4.6 tonnes of gold. This paper describes these outstanding achievements. INTRODUCTION In 1809 the Kingdom of Sweden was defeated by the Russian Empire in the Finnish War, and Finland was annexed by the Russian Empire as the autonomous Grand Duchy of Finland. Finland’s commercial and industrial interests were developed to serve the Empire’s needs. Finland had raw materials, but at the start of the 20th century Finnish industry was based on the use of wood as raw material and agricultural products. The mining industry was only a few percent of the country’s industrial production.

Jan 1, 2019