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By S. Tshikhudo

This paper outlines an investigation into the recovery of saleable fractions of coal from ?as-arising? South Export Plant effluent streams, using Baleen microscreening technology. South Export Plant, a subdivision of BHP Billiton Energy Coal SA (BECSA) Coal Processing, is a two-module plant treating 2000 t/h. The nominally -150 µm coal is untreated and is therefore passed from classifying cyclones to the thickeners for process water recovery. The thickened underflow is pumped into a series of slurry cells for further settling and recovery of supernatant water. The marginal quality, moisture content, and handlebility of this settled material renders it unsuitable for inclusion into saleable products and it is thus stockpiled and trucked to designated pits for disposal. Over the years, stockpiling and trucking has become an overly expensive exercise. In an effort to recover some of this cost, a task team was assigned to investigate less costly options to process slurry across BECSA plants. Various technologies such as froth flotation, sieve bends, and Reflux Classifier were considered, although the results were generally not beneficial ? this could be attributed to weathered/oxidized coal. A decision was made to pursue an alternative approach by testing the suitability of the new ?Baleen Filter?. The concept is to screen out the higher-grade fraction (+50 µm) as saleable product and reject the finer fractions to the slurry ponds. The Baleen Filter was found to effectively screen at an acceptable efficiency between 94% and 99.99%, with a very sharp cut-point (d50 and Ep). The actual yields from the screening results were better than the predicted yields in terms of both mass and energy as predicted from feedstock analysis.

Jan 1, 2014

By George C. Lipsey

FOR many years the mining of coal and other minerals occurring in sedimentary formation has been planned according to information obtained by diamond drilling. Placer mines have used churn drills for sampling and estimating deposits, and some metal mines have used the method either wholly or in part. The diamond drill method, therefore, although not entirely new, has come into more prominence in the last few years. Several large-scale developments now being planned or in the production stage have used this method in varying degree. This paper will describe the operations of the Howe Sound Exploration Co., Ltd. at Snow Lake, Manitoba, where a gold mine was developed and equipped for mining and milling wholly on the basis of diamond drilling results. The property has been in operation nearly 3 years, so that results can be compared with the estimates and assumptions made from the diamond drilling. In developing the subject it is necessary to outline as briefly as possible the normal method of mining, from discovery to production stage. Table I sets forth these various stages, the methods of financing, the percentage of total cost, and the relative time interval required for each section of the work. The plan presented in this table cannot be applied to all mining developments, but it gives a basis on which to work. In Stages 2 and 3 the investment is moderate, and if the property does not respond favorably the options or interests can be terminated. Stage 4 represents a sizeable investment, and it is more difficult to drop the property after this part of the operation is completed. Stages 2, 3, and 4 may be repeated several times by different groups or companies before Stage 5 is reached, but when Stage 5 has been started the total capital investment must be provided and there can be no turning back. It will be noted that there is a total of 116 units of investment in the overall cost if one company proceeds with Stages 2, 3, 4, and 5 in order. Stage 4, however, consumes time which is important when markets are being considered. Bringing a property into production when metal prices are low is much more difficult than when the prices are high, and if this stage can be set aside production is attained more quickly. When Stage 4 is eliminated the total becomes 86 units of cost for Stages 2, 3, and 5, but an adjustment of 14 units must be added to Stage 5. These 14 units represent the useable value of Stage 4. Thus Stage 4 represents an investment of 30 units if carried on before Stage 5, but 14 units if included with Stage 5. This is accounted for by the fact that development work can be accomplished more cheaply under Stage 5, since everything that is done is planned on a production basis with permanent services. Stage 4 must of necessity be governed by expenditures, and a great many temporary services, including housing and power plants, must be provided. The prospect shaft is a small one, and drifts,

Jan 1, 1952

By A. J. de Witte

INTRODUCTION The importance of determining dips of subsurface formations in wells, especially in wildcat wells, is self-evident and need not be enlarged upon. Various dipmeter instruments have been described in technical literature1,2,3,4 over the years. The introduction of the continuous dip-meters3,4 has made the earlier types obsolete. Continuous dipmeter logs are run by at least two different major service companies. Included as part of the service are a limited number of dip calculations—usually 10—made at different levels from the log. Very frequently this is the total result of a dip survey. It is believed that dipmeter surveys would be of considerably more advantage if geologists and engineers were able to make their own calculations from the original records. Determination of dips from the dipmeter record starts with the correlation of three log curves taken at different sides of the borehole. It cannot be over-emphasized that the quality of the results hinges primarily on this correlation. Nobody should be better qualified to correlate logs than a subsurface geologist or field engineer. The next requirement is to have an easy yet accurate way of computing the dip from the correlated points. Service companies use a mechanical device for this purpose. Directions for a graphical method were given in a recent article by B. 0. Prescotts. This paper describes a different graphical technique which, in our experience, is simpler, faster and at least as accurate. The entire manipulation is carried out on the stereonet in a number of easy moves. The paper is divided into two parts. This first part contains the introductory sections and full explanation of the geometry involved. The second part reviews the reading of data from an actual field log and summarizes the practical rules for computation of dips, the whole process being illustrated by an example. Following the outlined procedure, one could calculate a good many levels in a relatively short time, thus being able in particular to check suspected anomalies such as fault zones or unconformities in as close detail as the log will permit. GENERAL EXPOSITION OF THE METHOD Prnciple of the Dipmeter The position of a plane in space is defined by the location of three points, not on a straight line. Referring to Fig. I, if a borehole traverses a bedding plane, we can determine the strike and dip of the bed by finding three points which are not in line along this plane. Such points as A, B, and C are obtained with the dipmeter. The device consists of three electrodes (or caliper arms)1, 2, and 3 spaced at 120 angles, as schematically indicated, whose readings are separately recorded (Curves 1, 11, 111, to the right). When electrode 1 passes the boundary of the bed, Curve I will kick, etc. From these curves, we know the relative distances of points A, B, C, measured along the axis of the hole. If A is at zero, then B and C are at levels d1 and d2 up the hole,

Jan 1, 1957

By B. Pesic

The recovery of silver and gallium from flue dust with acidified thiourea solutions was examined. The studied parameters were temperature, particle size, thiourea concentration, acid concentration, and the presence of oxidants. The study of the extraction of these two metals was undertaken in sulfuric acid medium. Silver dissolution was dependent on all examined parameters, whereas gallium extraction was only dependent on the initial acid concentration. The experimental results suggest a new very efficient method for recovery of both silver and gallium. Almost complete recovery of silver and gallium can be accomplished within 90 minutes. A technological scheme for recovery of both metals is also proposed.

Jan 1, 1986

By P. A. Crane, N. Shirley, Aung Moe, H. M. Nicholson, D. Kan

Kazakhstan’s first commercial copper-leach solvent extraction-electrowinning plant, Central Asia Metals’ Kounrad project, was commissioned in April 2012. Exploiting low grade waste stockpiles produced during Soviet era open pit mining, a 10,000 tpa copper cathode plant was designed, constructed and commissioned within 18 months. Overcoming initial challenges with a workforce unfamiliar with the technology and with extensive periods of sub-zero conditions during winter the project has consistently met targets and produced high quality copper cathodes at cash costs of less than US$1/lb. Challenges encountered during the first two years of operation are described and future plans briefly discussed.

Jan 1, 2014

The negative charges on diaphragms of quartz, tungstic oxides, stannic acid, acid dyestuffs, soaps, and glass have for a number of years been explained on the basis of chemical equilibria—a hydrogen ion or light metal ion passes into solution leaving the solid complex ion negatively charged. While there are a number of observations which indicate a proportionality between the ( potential and the streaming potential it should be recognized that all the experimental evidence thus far presented shows that if hydrogen ion concentration is kept at a constant value, electrokinetic phenomena are influenced by salts. The multivalent ions exert the greatest effect on the reduction of osmotic flow. In the interpretation of all electrokinetic data the assumption is made that the potential of the solid remains constant. That this is not true is evidenced by the values of [ potential for quartz in contact with conductivity water as reported by the several investigators and by the authors of this paper. It is apparent that the nonreproducibility of surfaces will always prevent the obtaining of exactly comparable data. At best the experimental technique described depends on a theory which is capable only of showing a trend—another quartz surface should exhibit similar but not identical potential curves. It seems entirely reasonable to assume that the negatively charged laurate ions should have little or no attraction for a negatively charged quartz surface. The weight of previous experience in selection of flotation activators points unerringly to the usefulness of the multivalent metallic ions. The primary reason why the divalent barium ion is not available to two negative charges on the quartz surface might easily be that of distance between such adjacent negative charges. Otherwise activation could not occur, and the quartz could not assume the remaining unit positive charge of the barium ion. For this case no electrical double layer theory seems necessary. Clearly this is, at the negatively charged quartz surface, a case of partial chemisorption of the cationically divalent barium ion followed by the anionic neutralization of the remaining cationic charge by the freely moving laurate ion. It is hoped that the authors will continue with this interesting technique and that their results will continue to indicate trendwise what is known about the

Jan 1, 1957

By A. S. Swanson

The Babb fault vein in Crittenden and Livingston Counties, Ky. (fig. 1), is one of an extensive series of such veins in the Western Kentucky and Southern Illinois fluorspar district. First operated in 1904, the production to May 1, 1944, was estimated at 143,000 tons of fluorite ore. There are five mines (fig. 2) on the Babb fault zone - Guill ,Kentucky Babb, Delhi-Babb, Eagle-Babb and Howard-Barnes. The Bureau of Mines developed four of these properties by core drilling from July 9, 1943, to May 1, 1944. During that period a total of 8,177.8 feet of diamond drilling was completed in 28 holes.

Jan 1, 1950

By Carl-Henrik Knudsen, Rolf Tore Arnesen, Eigil R. Iversen

Recovery of iron and copper from acid mine drainage was studied in pilot scale at a continuous treatment plant at the mine site of the King's mine in Roeros, Norway. The average load on the plant was 1 m3/h. The process tested in this project is a three-step process consisting of: 1) Fe(OH)3 precipitation at pH 3.5, dewatering of iron sludge and transport to ferric chloride plant; 2) after filtration precipitation of residual metals at pH 9, concentration of metal sludge by flotation; 3) extraction of metals from the sludge with sulphuric acid followed by recovery of copper by electrolysis. The copper could be used for local production of different crafts, but this condition needs a further evaluation. The process economy at a full scale plant is evaluated.

Jan 1, 2000

By B. J. Arnold

Utility coal specifications usually contain a requirement for a maximum moisture content in order to prevent coal handling problems and to minimize evaporative losses in the boiler. As most coal cleaning processes are water-based, dewatering and sometimes drying steps are required to meet moisture specifications. Screens are often used for dewatering, with centrifuges and filters used to further reduce moisture in many cases. Algorithms to predict the performance of screens were developed for the Coal Cleaning Simulator running under Aspen Plus during a project sponsored by the U.S. Department of Energy and the Electric Power Research Institute. These algorithms include calculations for the size separation and a calculation for moisture content of the oversize product based on particle size and specific gravity distribution. This paper discusses the algorithm for predicting the moisture content of coal dewatered by screens.

Jan 1, 1997

By John W. Rozelle, David S. Bolin

INTRODUCTION During the normal course of a mining project, a feasibility study will be prepared that addresses all the aspects of the project including geology and mineral inventories, mining and geo- technical, metallurgical processing, infrastructure, environmental considerations and economic analyses. These studies are prepared for various reasons. Occasionally, a feasibility study will be prepared solely for use within a company as justification for development expenditures, however, these documents are frequently submitted outside the company either to raise funds from financial institutions or from public financing on one of the various securities exchanges. Mineral industry consultants are often called upon to participate in the preparation of feasibility studies or to conduct an independent audit and review of an internally prepared study. As the recoverable mineral resources form the foundation upon which the balance of a feasibility study is based, including the economic analysis of the project, the resource aspect of the study is examined very carefully by financial and securities organizations. The balance of this discussion is directed toward the various geological and resource estimation aspects of feasibility studies that experience has shown to be of particular importance. Much of this discussion will be directed toward open pit gold operations, although the principles are applicable to any type of deposit. 2.0 GEOLOGIC RESPONSIBILITIES IN FEASIBILITY STUDIES The principal geological effort in the preparation of a feasibility study is generally dedicated to the following areas: - Preparation of Geologic Reserve/Resource Estimates - Condemnation of Sites for Facilities, Access Roads, Leach Pads, Waste Dumps and Tailings Disposal - Collection of Geotechnical Data - Provision of Mineralogical and Petrological Data Pertinent to Metallurgical Processing - Selection of Appropriate Bulk Sampling Sites for Collection of Metallurgical Samples Each of these items will be addressed in detail in the following sections. 3.0 RESERVE/RESOURCE ESTIMATION An ore reserve estimate consists of the determination of a volume, or tonnage, and a grade for the deposit. There are numerous cases in the mining industry of projects that have experienced severe difficulties when operational ore grade, tonnage or metallurgical amenability have proven to be deficient compared to predicted estimates. An error in estimation of tonnage generally leads to more long term problems affecting mine life and production schedules. However, a significant error in the estimation of grade, particularly if mined grade is lower than anticipated, can have an immediate, and often disastrous, effect upon a project. The geologic preparation leading to an ore reserve estimate generally follows a logical sequence familiar to

Jan 1, 1987

By James R. Hein

Manganese nodules form at the sediment surface of abyssal plains throughout the global ocean at water depths of about 4000-6500 m. The nodules form a continuum from purely hydrogenetic (all metals from seawater) to purely diagenetic (all metals from sediment pore waters), which will determine which metals of potential economic interest are enriched in the nodules. Another important factor in metal enrichment is whether diagenesis in the bottom sediment is oxic or suboxic. There are several nodule fields in the deep ocean that have received much attention with regards to economic potential, the most important being the Clarion-Clipperton zone (CCZ) in the NE Pacific, where the International Seabed Authority (ISA) has signed 12 contracts for exploration areas, each spanning 75,000 square kilometers. Those nodules are rich in nickel (mean 1.3 wt. %), copper (mean 1.1 %), and manganese (28%), with moderate amounts of cobalt (mean 0.2%), molybdenum (587 g/t), and lithium (129 g/t). A conservative estimate is that there exists 21,100 million dry tonnes of nodules in the CCZ, yielding more manganese, twice the nickel, three times the cobalt, and four times the yttrium than the entire land-based reserve base (economic plus marginally economic plus subeconomic reserves) for each metal. Total rare-earth elements (REEs) as oxides average 0.09% in CCZ nodules, a potential byproduct even at that grade because the heavy REE complement averages 26%, compared to less than 1% for the large land-based carbonatite REE deposits. The price for HREEs is much greater than for the light REEs. For comparison, the REEs in deep-ocean ferromanganese crusts are 3 to 5 times greater than nodules and the HREE complement averages 18%. Of the 12 CCZ contracts, 7 are with States (China, France, Germany, Japan, Korea, Russia, and a group of States (IOM); and 5 are with companies: G-Tec Sea Mineral Resources NV, Nauru Ocean Resources Inc. (NORI); Tonga Offshore Mining Ltd (TOML), UK Seabed Resources Ltd (Lockheed UK), and Marawa Research and Exploration Ltd. Three other nodule fields that have received attention are in the Peru Basin, which has nodules very rich in manganese (34%) and lithium (mean 311 g/t), the Penrhyn Basin (Cook Islands Exclusive Economic Zone), with nodules very rich in cobalt (mean 0.4%); and the Central Indian Ocean Basin (CIOB), where India has an exploration contract with the ISA. CIPB nodules have mean nickel, copper, and manganese contents of 1.1%, 1.0%, and 24% respectively.

May 1, 2013

By J. P. Rider

The number of extended cut mining systems has increased significantly in coal mines over the last decade, and there is a constant push to extend the face ventilation systems to even greater setback distances. The need to maintain airflow at setback distances of12.2 m (40 ft) and beyond requires that methods for delivering intake air to the face be improved, or alternative ventilation techniques be implemented. The emphasis of this research effort was to examine various techniques to improve the distribution of available air at the mining face in order to reduce face methane levels. A study was conducted comparing the performance of a jet fan, used in conjunction with a machine mounted scrubber, with blowing curtain ventilation in a 27.5 m (90 ft) heading. In addition, a second study was performed on blowing curtain systems examining how scrubber flow, intake air velocity, and curtain setback influence face methane concentrations. Results showed that when the heading advances to a depth of 27.5 m (90 ft) blowing curtain systems were more efficient at diluting face methane levels when compared to a jet fan. Also, increases in scrubber flows and curtain velocity resulted in lower face methane concentrations for blowing curtain systems at 9.2 and 15.2 m (30 and 50 ft) setbacks.

Jan 1, 1997

By Robin Phinney

Certain statements in this presentation may constitute "forward-looking" statements which involve known and unknown risks, uncertainties and other factors which may cause the actual results, pe1formance or achievements of the Corporation, or industry results, to be materially different from any future results, performance or achievements expressed or implied by such forward-looking statements. When used in this presentation, such statements use such words as "may", "would", "could", "will", "intend", "expect", "believe", "plan", "anticipate", "estimate" and other similar terminology. These statements reflect the Corporation's current expectations regarding future events and operating performance and speak only as of the date of this presentation. Forward-looking statements involve significant risks and uncertainties, should not be read as guarantees of future performance or results, and will not necessa1ily be accurate indications of whether or not such results will be achieved. A number of factors could cause actual results to differ materially from the results discussed in the forward looking statements including, but not limited to, the factors discussed in the Corporation's annual information form filed on March 9, 2011. Actual results could differ materially from those currently anticipated due to a number of factors and risks. These include, but are not limited to, the failure to obtain necessary regulatory approvals, including environmental approval, the failure to secure full project financing, the failure to implement the recommendations set out in the Technical Report, the failure to raise financing to implement the recommendations in the Technical Report, risks associated with the mining industry in general (e.g., operational risks in development, exploration and production; delays or changes in plans with respect to exploration or development projects or capital expenditures; the uncertainty of estimates and projections relating to production, reserves and resources, costs and expenses; and health, safety and environmental risks), decrease in potash prices and exchange rate fluctuations. Although the forward-looking statements contained in this presentation are based upon what management of the Corporation believes are reasonable assumptions, the Corporation cannot assure shareholders that actual results will be consistent with these fonvard-looking statements. These fon\?ard-looking statements are made as of the date of this presentation and are expressly qualified in their entirety by this cautionary statement. Subject to applicable securities laws, the Corporation assumes no obligation to update or revise them to reflect new events or circumstances. This presentation is a market update based on the Corporation's public disclosure and does not form a part of any offer of securities, or constitute a solicitation of any offer to purchase or subscribe for securities. Any purchase of securities in the company should be made solely on the basis of information contained in the Corporation's public record which is available on www.sedar.com .

May 1, 2013

By C. R. Ramachandra, C. C. Patel

Flotation of chalcopyrite from a low grade ore was studied by using different xanthates and dixanthogens as collectors and by conditioning the flotation pulp with oxidizing gaseous systems. The improvement in the recovery of chalcopyrite with the oxidizing systems was mainly due to the auto-activation of chalcopyrite, leading to the enhanced adsorption of xanthates and dixanthogens even when the latter were formed by oxidation of the xanthates. With air or oxygen as oxidizing system, the amount of ethyl xanthate required at pH 8.5 was much less than normally employed in the industry, the recovery being 97-99% of chalcopyrite with a concentrate of 24% copper. With isopropyl and isoamyl xanthates, the recovery was improved but the grade was lowered. By conditioning the pulp with air-CO2 or oxygen-CO2 mixture, the recovery was lowered with ethyl xanthate only, mainly due to the depressant action of bicarbonate formed. Flotation with dixanthogens gave cleaner concentrates of chalcopyrite. because iron pyrite did not float with these collectors. Of the dixanthogens tried, diisopropyl dixanthogen gave the highest recovery, 99% chalcopyrite with a concentrate containing 28.5 to 30% copper, when the pulp at pH 8.5 was condi-tioned with air. Isopropyl dixanthogen was found to be the best collector of all the xanthates and dixanthogens studied. It was found that the adsorbed ethyl, isopropyl and isoamyl xanthates at chalcopyrite surface, activated by cupric sulfate, and at iron pyrite surface were oxidized to the corresponding dixanthogens to different degrees in presence of air, oxygen, air-carbon dioxide mixture and oxygen-carbon dioxide mixture.' The dixanthogen formed at chalcopyrite surface was, however, firmly adsorbed while that at iron pyrite surface was easily desorbed. Since dixanthogens formed have greater hydrophobicity than the corresponding xanthates, as revealed by contact angles,1,2 the former were expected to improve the flotation results. It was, therefore, thought desirable to find out the effect of air, oxygen, air-carbon dioxide mixture and oxygen-carbon dioxide mixture on the selective flotation of chalcopyrite when individual xanthates were used as collectors. The effect of direct use of dixanthogens as collectors on the selective flotation of chalcopyrite was also investigated. For these studies, low grade chalcopyrite ore, received from the Indian Copper Corp., Ltd. (I.C.C.), Ghatsila, Bihar, was employed. At the milling plant of the I.C.C., chalcopyrite is floated at a pH of 8.5, employing ethyl xanthate as collector. EXPERIMENTAL Materials Employed: XANTHATES - Ethyl, isopropyl and isoamyl xanthates of potassium were prepared in a pure state by the methods described earlier.4 A 0.03% aqueous solution of a xanthate was freshly prepared in air-free distilled water for flotation studies. DIXANTHOGENS - Diethyl, diisopropyl and diisoamyl dixanthogens were prepared by the oxidation of xanthates by iodine and purified by extraction with petroleum ether. A 0.03% alcoholic solution of dixanthogen was used as a collector. FROTHER - A 0.2% solution of pine oil in ethyl alcohol was employed. SODIUM CARBONATE - A 0.4 N solution was used. ORE - The composition of the -100 +200-mesh size (Tyler standard) powdered ore, as determined by standard methods of analysis, was: copper 1.57%, iron 10.29%, sulfur 3.15% and acid insolubles 66.10%. Iron present as chalcopyrite (CuFeS2) was 1.38%, while that present as iron pyrite (FeS,) was 1.37%. Both chalcopyrite and iron pyrite were found to be liberated from the gangue minerals and from each other, as observed under a microscope. Flotation of Chalcopyrite with Xanthates in Presence of Different Gaseous Systems: Fifteen grams of -100 + 200-mesh size ore were shaken well with 100 ml of distilled water. The pH of the pulp was adjusted to 8.5 by the addition of sodium carbonate solution. Then 0.1 ml of xanthate solution (0.005 lb xanthate per ton of the ore) was mixed with the pulp, and the latter together with washings transferred to a

Jan 1, 1963

By G. Townsend Harley

STOPING methods in which shoveling plays an important part are gradually being replaced by other and cheaper methods. But there will always be considerable shoveling done underground in stopes as well as in drifts, tunnels, urinzes, and shafts. At the mines of the PhelpsDodge Corporation at Tyrone, N. M., the cost of shoveling in all stopes in 1917 amounted to 24c. per T. In the top-slice stopes for the same period, it cost 27 c. per T. or 16 per cent, of the total cost of these stopes. The tonnage for shovelers from all stoping was 9.3 T. per man, and for topslicing 8.2 T. per man. These stopes were not unduly hot, and there was not more than the usual amount of timber to interfere with the work of the men. The tonnages obtained per shoveler were considered low; first, because of a poor grade of Mexican labor, many of the men having come in from railroad grading camps; and second, because of a poor spacing of raises, especially in the top-slice stopes, where, in general, they were spaced 25 ft. by 66 ft. (7.6 by 20.1 m.) centers. The average wage per laborer shift was $2.67 during the year. It was thought, however, that even under these conditions the men were not producing the tonnage that they should, so, with the consent of the management, the writer undertook to determine how the general efficiency of the underground shoveling could be improved. No predetermined plan for conducting these experiments was arranged because we had no definite ideas as to the scope of the work or the number of elements into which the investigation would resolve itself, before all of its phases could be determined. We were sure, however, that any work that would thoroughly cover the ground would have to be in the nature of a systematic time study, combined with a course of instruction in correct shoveling methods and adequate and intelligent supervision of the work. Two or three companies in the Southwest have done some work to determine the proper shovel to be used undcrground, but so far as is known the work has been limited to equipping certain parts of their mines with a particular type and size of shovel and thereafter watching the cost and efficiency records. In each case it seems to have been the shovel that held the 21-lb. (9.5-kg.) load that gave the bcst results. Excepting personal communications from these companies, the only data available on scientific shoveling are contained in F. W. Taylor's book, "The Principles of Scientific Management," and D. J. Haucr's article, in The Contractor for March 29, 1918, "A Hundred Hints for Shovelers." This paper discusses and draws conclusions from several thousand time-study readings, taken both on the surface and in the mines for nearly a year. A sufficient number of readings were taken on each factor in the problem for the plotting of curves of the performances and to obtain accurate indices of the work to be expected from this class of labor. The work is not as complete in all of its details as we would like to havc it, because we were forced to stop the work temporarily, owing to the numbcr of men going into the National service and our inability to get others who could make time studies. The results obtained so far, however, have been of such a startling nature that we have decided to submit them at this time, subject to future modification. It is hoped, also, that a free discussion of this paper will lead to a disclosure of any errors that may have becn embodied and offer some valuable suggestions for the conduct of future work.

Jan 1, 1920

By John S. Brown

IN 1925 the writer undertook a detailed statistical study of all producing areas in the Joplin district as a basis for evaluating programs and measuring objectives. For this purpose, the published figures in the yearly volumes of Mineral Resources were used, supplemented for earlier years by publications of the Missouri Geological Survey and other local and less official sources. When all else failed, the available data were projected backward to hazard a reasonable guess as to the unrecorded early output of important areas. Fortunately, the proportion of such prehistory production is not a large factor in any of the totals. These results were used during the next few years to measure the relative importance of various producing areas and to predict the peak period of development of the all-important Picher field. For the purpose of this review, the charts have been completed to the end of 1950. During World War 11, the U. S. Bureau of Mines became interested in a similar study and issued comprehensive statistical tabulations of data up to 1945 ( Info. Circular 7383), which have been checked against the figures used herein. This tabulation, however, does not include all the earlier data used by the writer nor does it offer any estimates of the wholly unrecorded era in the beginnings of the earlier camps. The area covered in this study is shown in Fig. 1 on which are indicated the relative location and approximate outlines of the principal producing camps. This also shows the approximate yield to date of each major camp in terms of combined lead and zinc concentrates. The output of zinc concentrates is roughly seven times that of lead. Hence, the economy of the district has depended primarily on the price of zinc, with lead as an important byproduct. Over much of the productive period, lead concentrates averaged about twice the value of zinc concentrates per ton, and in certain mines or areas the proportion of lead to zinc was substantially above average. The Joplin district is largely flat prairie but is partly moderately dissected, partially wooded land with a relief generally less than 100 ft. The rocks are almost flat-lying, nearly parallel to the surface, and the chief ore formation is the Mississippian Boone limestone, including its cherty phases. This formation either outcrops in the producing areas or is covered by a thin veneer of Pennsylvanian shales. Virtually all the ore occurs within 400 ft of the surface, and a large part at less than 300 ft in depth. Most of the land was divided into small farms or town lots before mineral development; tracts seldom exceeded 160 acres, and averaged considerably less. Mineral rights followed the surface ownership, segregation was rare, and a system of leasing for mineral development became well established early in the region's history, many landowners deriving small to sizable fortunes from royalties. Because of the shal-lowness of the ore and other factors, prospecting and mining was cheaper than in almost any comparable mining district in the United States. This situation, coupled with the widely divided land ownership, offered a fertile field for promoters and speculators and led to the rise of many small mining concerns. Only in its later history, under stern economic compulsion, has control tended to centralize in a few companies. Under these conditions, any important new discovery or successful development had much the effect of a gold rush or an oil boom. Every property in the area was leased quickly, promptly drilled, and, if ore was found, it was soon on the market. Many companies and individuals participated, and the average producing lease-hold probably was about 40 acres in extent. Any important field thus was attacked by anywhere from 10 to 100 or more producers. Production zoomed, eventually steadied or wavered, and ultimately subsided, leaving a desolation of tailings mountains, cave-ins, empty housing, and wreckage. The object of this paper is to depict the pattern of this process, so far as metal production is concerned, and to note the way in which it reacted to economic and political pressures. Production Charts In Fig. 2 is charted the production record, in tons of lead and zinc concentrates combined, of eight of the principal camps, which together account for approximately 99 pct of the total district production, over the years from 1870 to 1950. This period covers all but the very minor beginning of mining history. Two important camps are divided by state lines; hence, it has been necessary to combine production records for the two portions, based on estimates that may be slightly in error. Certain camps are sub-dividable into important units for which separate figures are available in whole or in part and have been charted as fractions of the major unit. The corresponding price of zinc is shown above all the charts. Three camps, Aurora, Neck City, and Galena, show a remarkably symmetrical graphic pattern, which is interpreted as the norm. The curves rise steeply to a peak, level off for an irregular interval, and then drop sharply to zero on a slope corresponding roughly to that covered by the initial rise. The three portions of these charts seem appropriately characterized by the designations of youth, maturity, and decline. On the whole, with some irregularities, the production in each of the three periods seems to be almost equal. A fourth camp, Granby, fails to conform to the normal pattern. It exhibits a very long period of reasonably uniform, stabilized production corresponding to maturity, followed by a rather precipitate decline. Its youth is hidden in the era of prehistory. This habit of steady, long-continued production at an even keel is attributable to the fact that this camp, more than any other, was controlled largely by a single principal owner at any given period over most of its history and this permitted the imposition

Jan 1, 1952

By R. D. Lord

The search for the best practical means of storing sulfide bearing tailings, where there is no residual excess of carbonate material is discussed in this paper• Usually the sulfide content decomposes, with the aid of bacterial action, and the resulting sulfuric acid escapes, along with any heavy-metal solutes, through embankments that are usually porous to some degree• The problem is typified in the tailings of the uranium operations of Elliot Lake, Ont., where mining started some 20 years ago• The approach to tailings disposal paralleled the practice for other hydrometallurgical plants treating gold and base-metal ores• Impoundment areas were designed to retain solids, and a clear and neutral overflow was considered satisfactory practice• Now experience has shown that these areas, some of which have been idle for over a dozen years, release acids in seepage and overflows to an unacceptable degree• To protect natural water courses, neutralizing plants are operated wherever required• Lime slurry is fed continuously into the tailings outflows in a quantity sufficient to raise the pH to 8•5 and precipitate heavy metals that may be in solution• The objection to this procedure is that the plants will require servicing indefinitely, unless a better remedy is found• The problem differs only slightly from that common to base-metal concentrators in that here the ore has been leached with sulfuric acid for the recovery of uranium• Any native content of calcareous material has been digested, and only that added for final neutralization is available to maintain a pH unfavorable to bacterial activity• Chemical oxidation slowly lowers the pH and when this reaches a level of 4•5 or less, bacteria become active and greatly accelerate the formation of acid. The bacterial process is probably at least ten times as fast as the chemical oxidation• Location and Processing The operations referred to, uranium and one copper mine, are located at approximately 46°N and 82°W longitude• This is typical Canadian Shield country, a land of lakes, deeply glaciated and rocky, with sparse soil which supports mixed forest cover• Drainage is to Lake Huron, 25 miles to the south• Average temperature is 45°F, ranging from -40° to +95°F• Annual precipitation is 38 in•, about half of which is snow• The ore is Precambrian, quartz-pebble conglomerate, with mineralization in the matrix• From 5 to 10% pyrite is present• All known means of pre-concentration have been tested, but a bulk sulfuric acid leach has proved the most efficient. Tailings have from the outset been neutralized before release• Current practice is to add ground limestone to bring the pH to 4•5, and then lime to raise the value to 10•5• Environmental regulations have recently been increased and the foregoing meets the new standards• Separate measures are taken to precipitate radium• Remedial Measures Since the outstanding environmental problem is the oxidation of pyrite by bacterial action, the solution is to contain the products, or arrest the process• Given the ambient temperature, favorable half of the time, four items are essential to the activity• 1) Pyrite• 2) Moisture pH < 4•5. 3) Oxygen• 4) Bacteria• Removing any one of these out of the range of tolerance will bring the reactions under control• A variety of proposals considered, and a number tested for the arrest of the process, are: (a) render embankments impermeable, (b) provide an impermeable cover, (c) cover with an oxygen absorbing layer, (d) provide a vegetative cover, (e) flood the site, (f) remove pyrite from current tailings, (g) add excess limestone to current tailings, (h) poison the bacteria• Bank Seal-On existing impoundment areas, where the embankments are several thousand yards in length, it is believed that any program of injecting sealants can have small chance of success• However, a moisture barrier is an indicated specification for future construction, and this can be highly expensive• Surface Seal-Depending on the configuration of the deposit, the downward travel of water should be prevented, and oxygen excluded• Burying a plastic membrane just below the surface has been considered, as has the application of a liquid sealant that would penetrate the surface. The objection to these remedies is the excessive cost of dealing with large areas and the expectation of only temporary benefit as a result• Frost penetration is over 4 ft, and frost action breaks up asphalt paving and all but heavy concrete in a few years• Organic Layer-An oxygen-absorbing layer, such as bark fines from paper mills has been proposed as a surface treatment• Cultivated into the tailings such material might be expected to arrest subsurface oxidation for some years• Estimates are 100 tons per acre of bark fines, or 35 tons per acre of sawdust, and these enormous quantities do not so far give assurance of providing a long-term remedy• Vegatative Cover-Several obvious benefits would result from a good growth of grass or other vegetation on abandoned tailings• While restoring the natural green of the tract the growth would prevent wind-blown dust and reduce erosion• Subsurface oxidation should be reduced, as well as the upward movement of ground moisture as occurs in dry weather. To this end, considerable research and field testing has been carried out to arrive at a formula - a prescription which will provide a self-sustaining growth on the tailings surface, or at least one that would survive with reasonable maintenance attention. Many test plots have been run with different combinations of surface treatment and seed mixtures. Generally, by addition and close cultivation of limestone, lime, and fertilizers, technical success has been demonstrated• Plants with a high tolerance for acid soil seem the more hardy, and a pH above 3 is indicated so that nutrients can be absorbed• Recommendations are for 12 to 15 tons of

Jan 1, 1977

By G. Guzman

In the framework of the optimization of Teniente Converter operation the new system of automatic flux feeding through gar-gun was implemented in the Potrerillos Smelter, Codelco Chile, Salvador Division. The control algoritm based on simple iron and silica balance adjusted by empirical coefficients was incorporated in the logical programmer. The programmer works in real time gathering information from the plant instrumentation events and introduced chemical assays of concentrate, slag and flux. Through the auto-tuning of the automatic operation the criterions of optimal flux dosage were determined. Four month of operation of the automatic flux feeding system demonstrated the stabilization of magnetite content in the slag in the narrow range from 18 to 20% and the decrease of flux consumption, corresponding to annual savings US$ 340,000. Additionally, the automatic system helps operators in the control room in fast determination of optimal fluxing and avoid the errors in flux dosage affecting the slag quality and its impact on refractory wear.

Jan 1, 2007

By Allen H. Rogers

THE need for coordinated effort on public problem by engineers has long been felt. Early in June there will assemble in Washington a conference composed of delegates from all the engineering organizations in the country. The call for this conference is issued by the Joint Conference , Committee of the four national societies as the completion of its work on the organization of a congress composed of delegates representing all the engineering organizations of the United States. The plan for this confederation was presented at a joint meeting of the administrative boards of the national founder societies on Jan. 23, 1920; and was unanimously approved in principle by that meeting. The work of the committee since that date has consisted in drafting a constitution of the organization to be presented at the conference for discussion and, after such amendments as may be adopted, acceptance.

Jan 1, 1920

By F. D. Wright

Production of a coal preparation plant is governed by many restrictions, such as the tonnage of different products and blends that can be sold within a given period, capacities and output proportions of the cleaning and sizing units, blending proportions, quality specifications, and costs and prices for the various products. Determination of the tonnage of each product and blend that should be made in order to obtain the maximum profit can be difficult unless a systematic method such as linear programming is used. In this paper the basic method of linear programming is described briefly. The Old Ben No. 9 preparation plant is used as an example to illustrate in detail how eqwations can be written to form a linear programming model of a coal preparation plant. Three sample problems, each requiring 56 or more equations and 63 structural variables, were solved with an IBM 650 computer. Linear programming is one of several mathematical tools used for operations research. It has been applied to many fields and has been used by a number of industries either to maximize the profits of certain operations or to minimize costs. P. B. Nalle and L. W. weeks&apos; have described the use of the method by the Riverside Cement Co. to minimize the cost of blending raw materials to make portland cement. Their problem is to obtain at minimum cost a mix with certain specifications from a number of possible materials which have various costs and various amounts of CaO, SiO2, Fe2O3, and other constituents. In a paper on the use of linear programming by the National Coal Board in England, K. B. Williams and K. B. Halley2 describe how the transportation method, a variation of linear programming, is used to minimize the cost of sending 37 grades of coal from 28 mines to seven central washing plants which produce coal for furnace coke and foundry coke. The various coals have different percentages of volatile matter, moisture, sulfur, ash, and phosphorous, so there is considerable choice in how they can be blended to meet the specifications of the two products. The purpose of this paper is to show how linear programming can be used to maximize the profits of a coal washing plant which produces individual final products as well as blends. The Old Ben Corp. furnished assumed sample data from their Old Ben Mine No. 9 preparation plant for this investigation. However, the data that have been used are entirely the author&apos;s responsibility. METHOD OF LINEAR PROGRAMMING Numerous articles and books have been written on the theory and applications of linear programming.3-5 However, since the method has not been widely applied by the mining industry, a brief, nontheo-retical discussion of its basic method seems to be in order. Linear programming (Table I) is used to determine the best possible solution to a number of interdependent activities. It is essentially a method for making systematic selections from a number of possible solutions to determine positively the optimum solution. There may be other equally good solutions but no better ones. Each activity must have linear coefficients and there must be a criterion to judge how good the solution actually is. Linear programming is different from the solution

Jan 1, 1961