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By S. Banisi, M. R. Yarahmadi, S. R. Mortazavi

Nowadays the development of automatic on-line sampling systems has enabled large amounts of data to be available to plant managers. These data are inevitably prone to random and sometimes systematic errors. Data reconciliation methods are used to deal with random errors. In these methods, the measured data are adjusted in accordance to their error, so that the mass conservation law is satisfied at the same time. Measured data may also contain systematic errors, using these data in mass balancing software will result in inaccurate or even irrational results. In this paper three methods for detecting systematic errors (global test, node constraint test and maximum power constraint test) were presented. A step by step method to locate the systematic errors was demonstrated. Then, the new version of Movazen©, a mass balancing software, was introduced and used to locate systematic errors in the measurements taken from the Sarcheshmeh pilot plant copper flotation circuit. Keywords: Mass balancing, data reconciliation, random error, systematic error

Jan 1, 2020

By Gonzalo Bernstein, David G. Meadows, German Naranjo, Edgardo Tapia

"During the last decade Antofagasta Minerals’ Los Pelambres Mine has dramatically increased its ore throughput following the fast initial phase ramp up through a series of planned incremental expansions. The original 85,000 mtpd, two SAG line plant was commissioned in 1999 and rapidly exceeded nameplate tonnage in only a few months. Following this, the inclusion of a pebble crushing circuit resulted in throughputs in excess of 114,000 mtpd and a recent, final third line SAG and ball mill addition have taken the plant to over 175,000 mtpd even with increasing ore hardness and changing ore feed characteristics. The progressive changes are described in this paper together with how the technical challenges have been met.INTRODUCTIONMinera Los Pelambres is a world class copper-molybdenum operation located in the Salamanca area in Chile, in region four, approximately 500 km north east of Santiago, the capital of Chile. In the late 1990s Antofagasta Minerals developed the Minera Los Pelambres project in Chile at an investment cost of approximately $1.3 billion USDs. The project construction was successfully led and completed by Bechtel Corporation in 1999 and within a matter of months the plant was already exceeding the nameplate capacity of 85,000 mtpd. In 2001 a pebble crushing circuit was added to raise the capacity to 114,000 mtpd. In the early part of 2003-2004, it was recognized that the hardness profile was changing with depth and that in order to maintain the plant capacity additional power would be needed in the SAG mills together with upgrading of the primary crusher. In conjunction with these changes a fifth ball mill and a second primary crusher station would also need to be installed. A further detailed engineering study and evaluation was initiated by Minera Los Pelambres. This was lead by Bechtel Corporation with major design contributions from FLSmidth and Siemens in order to establish the re-powering requirements for the SAG mills. These re-powering changes together with the inclusion of a pebble crushing circuit took the plant to up to 125,000 mtpd even with harder more competent ores. Finally in the last two years an additional grinding line has been added to take the capacity to consistently over 175,000 mtpd. The third line was installed in a record time of 256 days. A schematic of the current comminution circuit, discussed in this paper, is shown in Figure 2."

Jan 1, 2012

By M. R. Geer

The cleaning of fine coal, particularly for metallurgical use, is receiving greater attention. An increase in the proportion of fine coal in the run-of-mine product has been created by changes in mining practice, and a concur-rent deterioration in the quality of the fines has occurred because of more widespread full-seam mining. These two factors, augmented by the depletion of premium beds of metallurgical coal in some areas and the resulting shift to "dirtier" beds, are providing the impetus for adoption of improved cleaning techniques. For example, froth flotation, which has been used for years in a few preparation plants, now is gaining more rapidly in acceptance. Similarly, the dense-medium cyclone, which is used widely in Europe for making difficult separations, recently has been introduced to the United States. Regardless of the cleaning process employed, the finer sizes of coal can be treated with maximum effectiveness only when they are removed from the run-of-mine product and treated separately. No present process can treat the full range of sizes together with complete effectiveness. Flotation is applicable only to the coal finer than a top size ranging between 14-and 48-mesh, and cyclones require removal of all material finer than 0.5 to 0.3 mm. from the feed. Thus, the trend toward more-effective cleaning of the fine coal imposes the attendant problem of fine screening. Whereas in the past there was little need to screen finer than about l/4-in., present developments indicate that fine screening will become increasingly prevalent.

Jan 1, 1960

RISKGATE is an interactive online risk management system developed by the Minerals Industry Safety and Health Centre (MISHC; University of Queensland) to assist the mining industry in implementing continual improvement in management of major unwanted events so maximising health and safety performance. A comprehensive body of knowledge has been compiled from diverse mining industry experts for risk management of tyres, collisions, fires, isolation, strata underground, ground control open cut, explosions, explosives, manual tasks and slips/trips/falls. This paper presents an overview of the RISKGATE body of knowledge captured specifically for the management of unwanted events that stem from the handling and use of explosives (drill and blast) within surface coal mining operations. The knowledge presented is divided into a series of initiating events (listed below) within which there are further degrees of classification based on design, operation, and management concerns. Initiating events: ? Loss of control of explosives and associated chemicals in storage on site ? Loss of control of explosives and associated chemicals during transport ? Loss of control of explosives and associated chemicals during manufacture ? Loss of control of explosives and associated chemicals during handling and blast operations ? Loss of control of explosives and associated chemicals during disposal The system guides the user in generating substantive and leading-edge controls to assist in the design, management and reporting of drill and blast operations to meet stringent internal and external requirements. From a broader industry perspective, RISKGATE provides an environment for knowledge capture and knowledge exchange to drive innovation and best practice in the identification, assessment and management of risk. By capturing operational knowledge from industry experts, RISKGATE provides a cumulative corporate memory at a time of high personnel turnover in the coal industry.

Aug 1, 2013

By V. S. Gischig, G. Preisig

Hydraulic stimulation of deep reservoirs is an essential procedure for developing engineered/enhanced geothermal systems (EGS), but also in other Geo-energy contexts. The goal is enhancing permeability to such a degree that fluid flow rates within the reservoir are sufficiently high for an economically productive reservoir. Two fracture stimulation mechanisms are distinguished that may occur concurrently during the hydraulic treatment: 1) During hydro-fracturing (HF) new tensile fractures are propagated from the borehole by means of a fluid pressure overcoming the minimal principle stress d3 plus the tensile strength T0 of intact rock. In case of a pre-existing fracture oriented normal to d 3 , the fracture can be opened by exceeding d 3 only. Fluid injection is performed via small packed intervals for creating a stack of hydraulic fractures. 2) During hydro-shearing (HS), over-pressure induces slip along pre-existing fractures that are favourably oriented in the stress field for reactivation in shear. Such stimulation is usually performed in a large packed interval or in an open borehole. Which mechanism occurs predominantly during stimulation depends on the rock mass structure and in-situ stress field, but also on the orientation of discontinuities intersecting the open-hole section and thus being pressurized. Both mechanisms also differ in how they affect fracture permeability. While permeability gained by HS is mostly irreversible due to rearrangement of asperity contacts accompanied with shear dilation, permeability enhanced during HF reduces nearly reversible after pressurization unless proppant is used to ensure permanent apertures. In this contribution, we critically discuss the pros and cons of both mechanisms regarding efficiency of enhancing permeability and the associated hazard of induced earthquakes based on literature review and numerical modelling. We review literature that reports on how much permeability required for productive EGS reservoirs and how much permanent permeability can be gained by stimulation. We also compile information regarding the degree of seismicity induced during HFor HS-dominated stimulation procedures, together with conceptual studies that reveal characteristics of seismicity associated with the two mechanisms. Many observations and models indicate that HF may have a higher tendency of being aseismic, while felt seismic events are usually associated with HS. Further, we present a hydro-mechanically coupled fracture flow model that investigates, if and to what degree the stimulation strategy can be designed such that one of the two mechanisms is evoked and dominates over the other. It shows that HS may dominate in presence of larger persistent fractures nearly optimallyoriented in the stress field, even if HF is attempted in short packed intervals. The model further demonstrates that stress transfer during HS promotes the development of permeability across an anisotropic layer of the stimulated rock instead of a large volume. Hence, the study sheds light onto the feasibility of creating productive EGS reservoirs in crystalline rock at several kilometres depths.

Jan 1, 2015

By Kilgour I

The installation of a final cleaning column at Ok Tedi Mining Limited was carried out to increase final concentrate grade in order to reduce concentrate handling, shipping and treatment charges and to reduce penalty payments. To fast-track the project a decision was made to install a full-size column on one of the two parallel circuits, designed on information readily available and utilising the experience of mill staff and an engineering consultant, rather than delay installation while a pilot scale program was conducted. A 3.66 m diameter by 12.8 m high column was installed to produce 40 t/h of final concentrate from a combination of rougher concentrate and recleaner concentrate. The column design and sparger system were by Cominco Engineering Services Limited.

Jan 1, 1991

By Wesley T. Holmes

In this final stage of a three-part investigation of the smelting of titania-bearing iron materials, the Bureau of Mines conducted semicontinuous smelting tests in a laboratory electric arc furnace on four concentrates and one are. The objective was to substantiate the significance of liquidus temperature measurements in predicting satisfactory slag compositions in electric furnace smelting of titaniferous magnetites to produce pig iron. Previously obtained liquidus temperature data were used in determining flux requirements. Self-fluxing pellets of either ore or concentrate plus fluxes and reductants were made to produce pig iron and slags containing 0.2 to 0.6 percent titanium. Individual constituents of the slags generally agreed within 5 percent with the value of the corresponding constituent of target composition. Liquidus temperature of furnace slags generally agreed within 50° C of the value obtained from corresponding nominal synthetic slag. The results of the three phases of the investigation may serve as guide-lines for determining flux requirements in smelting a wide variety of titaniferous magnetites and other titania and iron-bearing materials.

Jan 1, 1969

By T. V. Lord

Minerals containing uranium have been successfully pre-concentrated by flotation at Queen's University, Kingston. In particular, uraninite, uranothorite, euxenite, and fergusonite, have responded to flotation to yield grades of 0.2 to 6.0 .per cent U .o. with recoveries of 85 to 98 per cent, from heads analysing approximately 0.10 per cent U3Os. Of the ten or so radioactive samples examined to date, two distinct groups were recognizable, firstly, those samples in which uranium was the sole constituent of economic interest, and, secondly, those in which Columbium (with minor amount of tantalum) was present in addition to uranium. For the purpose of this paper, the results of investigation on two samples of each of the above groups have been selected as typical.

Jan 1, 1945

By Gautam Chandra Mondal

An attempt has been made to carry out qualitative assessment of mine water discharges from the different mines of West Bokaro Coalfields of Hazaribag district of Jharkhand, India. A total of 33 mine water samples were analyzed for pH, electrical conductivity (EC), major cations (Ca, Mg, Na, K), major anions (HCO3, SO4, Cl, NO3), and heavy metals (Fe, Mn, Cu, Zn, Pb, Cr, Ni, Co, As). The hydrochemical data has been evaluated in terms of major ion chemistry, solute acquisition processes and suitability of mine water for domestic, industrial and irrigation uses. The pH of the analyzed mine waters varies from 5.27?8.43 indicating acidic to alkaline nature of the water. The electrical conductivity (EC) values varied from 276 µScm-1 to 1598 µScm-1 for measured samples and spatial differences between the EC values reflects the wide variation in lithology, surface activities and hydrological regime prevailing in the region. The anion chemistry of the analyzed water samples shows the dominance of SO4 and HCO3 with a minor contribution from Cl, NO3 and F. The observed high values of sulphate in the mine water of the area may be attributed to the oxidative weathering of pyrites. On average chloride is contributing 5% of the total anionic balance (TZ-). Fluoride and nitrate have low concentrations and on average contributing <2% the total anions. The cation chemistry indicates that 79% of the samples are Ca>Mg>Na>K, while the 6% belong to Ca>Na>Mg>K and 1% of Ca>Mg>K>Na. Weathering of rock forming minerals and cation exchange processes normally control the levels of these cations. Concentrations of some heavy metals (Fe, Mn, Ni, Cr) are found to be well above the prescribed permissible levels recommended for the drinking water. The water quality assessment for drinking purposes indicates that most of the mine water having high TDS, total hardness (TH) and sulphate concentrations. This suggests that mine water of the study area is not suitable for direct use in drinking and domestic purposes and needs a proper treatment before its utilization. The assessment of water for irrigation uses show that the water is good to permissible quality and can be used for irrigation uses. However, high values of salinity, %Na, SAR, RSC and Mg-hazard at certain sites restrict its suitability for agriculture uses. Keywords: mine water; ion chemistry; quality assessment; West Bokaro coalfields

Sep 1, 2012

The Woodlawn base metal sulphide deposit, located 50km S of Goulbum, New South Wales, was mined as an open pit operation for 10 years from 1977, producing 8 million tonnes of ore. Underground development began in 1985 and reached full production of 500,000t/yr as open pit mining ceased in mid 1987. Underground development highlighted the extensive deformation and intense alteration associated with the deposit and the host rocks.The deposit occurs above an acid volcanic pile in a sequence of interbedded and interfingering shales, cherts, volcanogenic sediments and pyroclastics of mid to late Silurian age intruded by dolerite sills.

Jan 1, 1990

By Roger T. Laine

"The Carswell structure, located in the western part of the Athabasca Basin, northern Saskatchewan, has been extensively explored since 1967. This exploration has been successful and lead to the discovery, in 1970, of the D sandstone-hosted unconformity related deposit. By the end of 1984, five additional basement-hosted unconformity related deposits had been delineated: the OP and N discovered in 1970, the Claude in 1971, the Dominique-Peter in 1981, and the Dominique-Janine in 1984. The D deposit has been mined out; the Claude open-pit has already produced more than 500 tonnes of uranium, the development of the Dominique-Peter underground mine (first phase) has been completed, and production from the latter started in 1985.All six deposits are located in the southern part of the Aphebian metamorphic core, 18 km in diameter, which has intruded at least 1250 metres of Helikian Athabasca Group sediments during a cryptoexplosion event at 480 Ma, unrelated to the emplacement of uranium mineralization. The metamorphic rocks, affected by the Hudsonian Orogeny, are divided in two lithostratrigraphic units: the Peter River gneiss, aluminous gneiss series, and the Earl River complex, mixed quartzofeldspathic and mafic gneiss series. A well developed regolith profile is recognizable in the south along the contact with the sandstone and inside the metamorphic core sometimes associated with conformable sandstone beds.Basement and sandstone-hosted uranium deposits show many similarities:-linkage to unconformity;-mineralogy and geochronology with a first phase mineralization characterized by a polymetallic sulphide-uraninite-brannerite assemblage dated at 1150 to 1050 Ma, a second phase characterized by a simple sulphide-uraninite assemblage dated at 890 to 820 Ma, and a third phase characterized by a pitchblende-hematite dated at 380 Ma; and,-linkage to early Hudsonian structures with remobilization of these structures during intense hydrothermal alteration resulting in Mg chlorite-illite halos."

Jan 1, 1986

By Faustino G. Prado

In the, quest for an economical and environmentally friendly reclamation of metallurgical dusts generated during pyrometallurgical operations, our goal has been the separation and recovery of most of the components, as well as the complete recycle of all the reagents. Effluents have been avoided, within acceptable economic parameters. This presentation, which can be applied to a large spectrum of dusts and other residual-materials-, has focused on the dusts generated by the operation of Secondary Copper Smelters. Those dusts show a wide range of components, including zinc, lead-, copper, cadmium, tin, chlorine and-alkalies. Our present work deals with the use of ammonium salts in the separation and recovery of-the above mentioned elements in a marketable form. Most of the chemistry and techniques involved have teen already confirmed in commercial scale operations. Others have gone through pilot scale tests or bench scale tests. The recycle of reagents has also been- confirmed at the same operating levels. A preliminary evaluation indicates that the chosen route, in addition to a viable economy, offers a complete and safe reclamation within the most strict environmental regulations.

Jan 1, 1995

By B. K. Mohapatra

Small manganese ore bodies in south Odisha, India occurs both as stratiform and stratabound ore bodies. These two ore bodies show contrasting mineralogical, petrographical and chemical characteristics. Stratiform type of Mn-ores consists of manganese, iron and silicate- phases in approximately same proportion. In contrast, stratabound ore bodies contain a high proportion of manganese oxides, silicates and phosphorus phases. Phosphorus is found within apatite in the stratabound ore bodies while in the stratiform ore it occurs as an adsorbed state with iron phase (goethite). Both ore bodies are low-grade (17-23% Mn) therefore do not have a market value. However, the stratabound ores are enriched in phosphorus (P2O5: >1%) compared to the stratiform ores. Both the ore types were processed through gravity, magnetic and roast reduction followed by magnetic separation techniques. The strata bound ore could be upgraded to 33% Mn from a feed of 23% by combination of dry and wet magnetic separation methods. But the phosphorus content also suits up (P2O5: 1.2%) in the processed product making it unsuitable for further use. The stratiform ores do not respond to either gravity or magnetic separation because of close association of manganese with iron phases. However, this could be beneficiated through roast reduction followed by magnetic separation. An ore with 17% Mn content could be upgraded to 40% Mn with a 42% yield and 78% recovery. The goethite, having phosphorus, converts to magnetite and is retained within the magnetic fraction, thereby reducing the concentration of phosphorus in the non-magnetic fraction, which becomes the product. This report demonstrates the influence of ore petrology in determining the up-gradational potential of manganese ore in the southern Odisha, India. Keywords: manganese ore, physical beneficiation, phosphorus rich ore

Sep 1, 2012

By Alice S. Allen, James Paone

SURFACE SUBSIDENCE IN MINING The consequences of subsidence are becoming progressively more serious as the consumption of an ever- increasing quantity of minerals conflicts with the needs of an expanding population for surface land area. The inevitable result will be an increase in the amount of land area on which undermining is restricted because surface subsidence must be prevented. Designation of land areas that may be undermined, provided that subsidence is controlled and kept within specified limits, also will increase. The decisions as to which land is to be used for urban, agricultural, or other purposes clearly involve considerations that go far beyond the scope of mining technology. Under the authority assigned to the US Bureau of Mines (USBM) by the Organic Act (May 16, 1910) and its succeeding amendments and pursuant regulations (30 U.S.C. 1-1 1 ), the Bureau conducts scientific and technologic investigations concerning mining and its related problems. Subsidence-control demonstrations that have been conducted under this authority include projects in Wyoming, West Virginia, Illinois, and the Pennsylvania anthracite region. The problenls for technology arc to devise methods for extracting minerals with controlled subsidence in varying geologic settings and for calculating or predicting the amount, extent, and characteristics of the subsidence that accompanies the extraction of our principal minerals mined by high-tonnage methods so that logical choices can be made from the possible alternative approaches. Where underground minerals and fuels are mined and removed, the voids that are created underground generate strong imbalanced stresses in the surrounding and overlying rock strata. The resulting readjustments in the rock masses may cause subsidence of the ground surface. Subsidence implies vertical collapse, because the most conspicuous component of movement is downward. However, the downward component is accompanied by differential horizontal strains that may be more damaging to man-made surface structure.; than the more apparent vertical displacements. SUBSIDENCE CONTROL The most widespread method of alleviating potential subsidence problems in undermined areas has been to backfill mine voids with mine refuse or some other in- expensive material that provides lateral support to the remaining mine pillars and vertical support to the mine roof and overburden. Most USBM backfilling work has been conducted jointly with the Pennsylvania Department of Environmental Resources in the anthracite region of northeastern Pennsylvania. These joint projects were conducted under the authority of the Anthracite Mine Drainage Control Act of July 15, 1955 (Public Law 84-162, as amended), and the Appalachian Regional Development Act of 1965 (Public Law 89-4. as amended). In addition to the Appalachian and mine-drainage projects, the Bureau has either conducted or participated in demonstration projects to develop the "pumped- slurry" method of backfilling mine voids. Three of these projects were conducted in Rock Springs, WY. The first was a field test of the pumped-slurry technique, conducted in 1970 under the combined participation of the City of Rock Springs, the US Department of Housing and Urban Development, the Dowell Div. of the Dow Chemical Co., and USBM (Candeub, Fleissig, 1971). The objective was to demonstrate that a large quantity of sand could be hydraulically injected under pressure through a single borehole, and that filling of the mine voids would be essentially complete. In the earlier "blind-flushing" methods, involving sluicing material through boreholes by gravity, quantities per injection hole ranged from 15 to 765 m3 (20 to 1000 cu yd), and the mine voids were only partially filled. In the first test of the pumped-slurry method, approximately 14 900 m (19,500 cu yd) of sand were injected success- fully through a single borehole. Subsequent information obtained through 43 monitoring boreholes indicated that mine voids below 11 330 m2 (2.8 acres) had been filled. As a result of the successful initial test of the pumped-slurry process, USBM conducted additional backfilling demonstration projects using the new technique. The first full-scale demonstration was carried out in the Green Ridge section of Scranton, PA, between 1972 and 1974. This project proved the feasibility of using the new hydraulic-injection technique to back- fill dry mine voids, as well as flooded voids, and it demonstrated that crushed anthracite refuse could be used as easily as sand in the process. Stabilization was pro- vided for about 202 300 m2 (50 acres) of Scranton, having a population of approximately 1000. Additional demonstration projects at Rock Springs, WY, between 1973 and 1975, resulted in the stabilization of about 364 200 m2 (90 acres). The work was necessary to preserve the physical and economic well- being of the city, because there have been numerous occurrences of subsidence during recent years. A demonstration project completed in 1976 in Rock Springs brought the total stabilized area in that city to 647 500 m2 (160 acres). At a cost of about $3,000,000, a population estimated at 7000 persons and property values exceeding $18,000,000 were protected. Presently. USBM is participating in subsidence- control projects involving backfilling in Pennsylvania, Illinois, and West Virginia. Eleven demonstration projects were in progress by USBM in 1977-eight in the Pennsylvania anthracite region and three in bituminous areas of West Virginia (one) and Illinois (two). The estimated total property value is over $100,500,000, and the total cost of the projects is expected to be nearly $20,000,000. The demonstration projects are designed to adapt the pumped-slurry technique to a variety of subsurface conditions, to increase efficiency, and to reduce costs.

Jan 1, 1982

By J. W. Town

This report summarizes the results of laboratory investigations conducted by the Federal Bureau of Mines to obtain hydrometallurgical information on the dissolution of mercury sulfides from ores, flotation concentrates, and pure mercuric sulfides. Statistical methods were used to design the experiments and evaluate the resulting data. Studies demonstrated that cinnabar is readily soluble in sodium sulfide, partly soluble in potassium sulfide, and essentially insoluble in ammonium sulfide. Additions of sodium hydroxide or potassium hydroxide to sodium sulfide solutions increased mercury extraction slightly; when either was added to potassium sulfide or ammonium sulfide solutions, mercury extraction was equal to that of sodium sulfide. Tests at the Albany Metallurgy Research Center showed that sodium sulfide would completely dissolve chemically pure mercuric sulfide, more than 95 percent of the cinnabar in 5-percent mercury flotation concentrates, and 60 to 90 percent of the cinnabar in low-grade ores. The double-leach tests showed that potassium sulfide would dissolve over 90 percent of the arsenic in mercury-arsenic products and 1 to 10 percent of the mercury; then sodium sulfide-hydroxide solutions would dissolve 80 to 90 percent of the remaining mercury and 5 to 10 percent of the arsenic.

Jan 1, 1961

By W. Chi, R. Regan, R. A. Washington

"In uranium mines, the purification of underground mine air can be accomplished by the removal of dust particles and radon daughters using filters. Full-scale statistically designed experiments with flow rates up to 10,000 cfm have been conducted at Denison Mines Ltd. on mine air filtration with vermiculite as a filter material. Various filtration bed thicknesses were chosen, with different face velocities through each bed thickness. Results on total dust, respirable dust and radon daughter collection efficiencies were analysed statistically. It is shown that these collection efficiencies are directly related to the filter bed thickness. However, the different face velocities seem to affect only the collection of radon daughters. Optimum collection efficiencies for a clean bed of about 30% were found for respirable dust and radon daughters, and about 41>% for total dust, using an 8-in.-thick bed and a face velocity of 500 fpm. Dust build-up on the filter bed during operation led to an increase in collection efficiency, reaching 40 to 60% after two weeks."

Jan 1, 1973

By Hinrichsen H. D, Rathburn D. R, Stieger W

The present situation in the field of iron ore pelletizing is briefly described with particular view of the Lurgi/Dravo travelling grate process. To meet todays pellet quality requirements with regard to physical and metal- lurgical properties it is often necessary to add basic fluxing materials. It has been de- monstrated that for certain ores, basic addi- tives provide a considerable improvement in fired pellet quality. The production of self- fluxing pellets is becoming increasingly attractive. One Lurgi/Dravo plant is now in operation and another is under construction which were designed for the production of self- fluxing pellets. Self-fluxing pellets may be considered as the primary or exclusive blast furnace burden and the installation of a sinter plant is not necessary if the pellet plant-can handle all in-plant fines. Reference is made to the first pellet plant utilizing iron ore concentrate, in-plant fines and basic flux. The use of solid fuel for pellet indura- tion is discussed because of the increasing prices and possible shortages of heavy oil and natural gas. Since there is a strong tendency to increasing plant capacities, the design fea- tures of travelling grate pelletizing plants for productions up to 6 million tonnes are Discussed.

Jan 1, 1974

By Martin H. Stanczyk

Two methods of recovering zinc from zinc ammonia complex solutions were investigated. One method studied was distillation of the ammonia from the solutions to recover a basic zinc sulfate, which after calcining to zinc oxide was reduced to metallic zinc using hydrogen at 900° C. The second method tested was electrolysis. The tests were made of synthetic and leach solutions analyzing about 80, 100, and 115 gpl (grams per liter) of Zn, NH3, and SO4, respectively. The leach solutions were prepared by oxidation leaching of a sphalerite flotation concentrate, assaying 55 pct zinc at elevated temperature and pressure. The distillation method recovered about 98 pct of the zinc in a product analyzing 97 pct zinc. Ammonia losses incurred during the leaching and distillation operations were about 0.1 lb per lb (pound per pound) of metallic zinc produced. Electrolysis of the solutions at a cathode current density of 25 amp per sq ft (ampere per square foot) yielded zinc deposits assaying in excess of 99.9 pct zinc. About 1.5 kwhr (kilowatt-hour) of power were consumed per pound of zinc deposited; the current efficiency was 72 pct. Other electrodeposition tests revealed that small amounts of ferric iron in the cell liquor were deleterious to electrolysis unless about 5 lb of glue per ton of zinc were present in the electrolyte. The use of glue and vigorous agitation of the electrolyte gave smooth zinc deposits from solutions containing as much as 0.1 gpl iron.

Jan 1, 1962

By Roger M. Buckeridge, Charles D. Reese, Silvio M. Martino, William T. Carey, Ansel H. Graham

FUNCTIONS OF RAIL HAULAGE SYSTEMS The principal application of rail haulage in under¬ground mining is in the movement of ore or coal pro¬duction from near the mining areas out to a processing plant. In some cases, such production haulage may end at a shaft pocket or slope belt conveyor transfer station, depending on the type of mine and the geo¬logical placement of the mineral being mined. Main-Line Haulage The basic reasons for choosing main-line rail haul¬age are high-tonnage production per shift and long haulage distances. The choice between rail and other systems, such as belt conveyors, is not easy and requires an integrated study of many factors: tonnage, distance, grade, product size, number of producing areas, ore reserve, safety, capital cost, operational and maintenance cost, and the mining system. Projecting a main-line rail haulage system for com¬parison with other systems requires a choice of many variables which are interrelated. The project must start with consideration of the mining system, tonnages per day, distances and grades, and loading and dumping conditions. Rail mine cars of different sizes and types are studied. High tonnage demands consideration of wider track gages. A suggested minimum for metal mines is 914.4 mm (36 in.); for coal mines, 1066.8 mm (42 in.). The choice of the train size should be governed by tonnage per shift, physical factors of loading and dumping, cycle time, and the number of trains that can be operated and safely controlled. After the size of the train is determined, the type, size, and specifi¬cations of the locomotive can be selected. The loco¬motive may be trolley, battery, or diesel; selection guides are discussed elsewhere in this chapter. The advantages of rail main-line haulage are high tonnage capacity, flexibility, reliability, safety, low operational and maintenance cost, and dual usage of the system (supply haulage, for example). The dis¬advantages are high capital investment, high installation cost, the requirement for more maintenance facilities, and restriction to certain ore reserves and mine pro¬jections. Supply Haulage Supply haulage is a basic requirement of any mine, but the type and volume of supplies vary with the type of mineral being mined, the mining methods, and the size of the mine. A mine developed for rail main¬line haulage gains supply haulage through the addition of a secondary track system to extend the main line closer to the producing areas. Additional haulage equipment, generally smaller, is also added. It is neces¬sary to design into the rail system more switch bypass points and train control systems. The track system standards demanded by rail main-line haulage con¬tribute to the reliability and efficiency of the supply haulage function, whereas a supply track system to supplement other haulage systems often suffers from poorer track and installation standards. When belt conveyors are chosen for production haulage, a rail supply haulage system is often installed rather than using rubber-tired vehicles for supply haul¬age. The conditions of the seam or ore body create these items for consideration: haulage grade of more than 6% favors rubber-tired vehicles; longer haulage distances favor rail; difficulty in maintaining a smooth roadway favors rail; and movement beyond the end of a rail system generally requires rubber-tired vehicles, making a single system rather than a dual system attractive. Although operation of a supply haulage system may be concentrated on a single shift of a multiple-shift mining operation, proper planning and supervision are required if the haulage is to be efficient. Often the newer personnel at a mine are assigned to supply haul¬age which may add to the difficulty of obtaining maxi¬mum efficiency. Supply haulage by rail offers the advantages of high reliability, lower maintenance, and high efficiency over long distances. The disadvantages are less flexibility, high installation costs, and rubber-tired vehicles for use beyond the inby end of track often required. Personnel Movement The efficient and comfortable movement of the miner to and from his work area has a significant effect on his daily output. Minimizing man-trip time also becomes more and more important as hourly wages increase. Personnel are generally moved over a rail system by one of three methods: a specialized personnel carrier to which a certain number of men who work in a certain area are assigned; a combination vehicle, used to carry personnel but also capable of being used as a light supply/utility locomotive during a working shift; and man-trip cars towed by a supply locomotive. The first and second methods are used in coal mines where long distances make speed and comfort important. The third method is more common in metal mines. Equipment Movement As mining equipment technology has advanced, many mining machines and their subcomponents have grown in size. This trend requires careful planning for adding specialized carriers to handle the movement of larger equipment. Continuous miners, longwall equipment, raise borers, and ventilation equipment are examples of machines which may be difficult to move and which require special handling. The selection of the weight, speed, and operating characteristics of supply locomotives used to move equipment is important. Larger mines may often justify a special group of loco¬motives for this application.

Jan 1, 1982

By Abel Aragón Espinoza

"El objeto del siguiente trabajo es describir los procesos implementados en la Planta de Cianuración del Sindicato Minero de Parcoy, S.A., situada en el campamento minero de Retamas, distrito de Parcoy, provincia de Pataz, La Libertad. La planta trata minerales piritososo con un contenido en oro que varía desde 6.5gr./tm. hasta 13 gr/tm., asi como los provenientes de otros minas como: Carlos Bernabé, Mishencanto, Cabana, Loros, Alaska. También desarrolla procesos como la trituración, donde el material triturado es depositado en una tolva metálica (120 T.M) para ser alimentado posteriormente por una molienda. Dicha máquina de molienda intermedia (New House de 5 Style B) sigue alimentando a través de una faja, para luego pasar a un molienda fina de 7 ' de diámetro por 10' de longitud, con alimentador de cuchara. Luego procede el tratamiento de las lamas, la precipitación y la fundición, donde el mineral es fundido en un horno de crisol abundante durante dos horas. Posteriormente la refinación comprende dos etapas: la refinación electrolítica en la que los ánodos son las planchas de grafito y los cátodos son planchas de grafito, convirtiendo al electrolito en una solución de ácido nítrico, con cobre y plata.y química, mientas que en la refinación química, el producto adquirido es cargado a los frascos Herlemmeyer y es atacado con una solución de una parte de ácido nítrico con dos de agua y una pequeña cantidad de alcohol. Como resultado final de este proceso, se considera al oro recuperado en un 92%."

Jul 20, 1956