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By Ryne Goodnough

INTRODUCTION The Wyodak Mine is located in the northern Powder River Basin of Wyoming and is thought to be the oldest continuously operating coal mine in the United States, having recorded yearly production since 1923. The mine has expanded over its 80-plus-year history from an underground room and pillar operation to a modem surface snip mine. Currently, the mine feeds four on-site power plants using conveyors, has a filly automated truck load-out, and ships coal by rail to nearby Wyoming power plants. Annual production of 4.8 MTFT is accomplished on a single-shift basis with 53 fill-time production, maintenance and administrative personnel. In 1990 a decision was made to modernize the coal feed system to the power plant complex with modular conveying equipment. The project involved constructing coal storage silos, overland conveyors, feeder breakers and modular conveyors to follow the mining benches. Total project cost in excess of $25 million was used to construe or commission: A 2,150-TPH 48-inch overland conveyor with horizontal curve A 2,150-TPH 48-inch plant feed conveyor 13 shuttle conveyors, each 150 feet in length 2 front-end loaders equipped with large coal buckets 2 crawler-mounted feeder breakers with compatible discharge conveyor 1 high-lift shuttle conveyor with adjustable discharge height 1 high-angle conveyor 2 4,500-ton coal silos with automated truck load- out capabilities. Conceptually, the plan was to develop 33 MTONS of reserve adjacent to the plant complex and use front-end loaders to tram ROM coal to in-pit feeder breakers. The primary crushing would reduce coal to 6-inch minus and feed through a series of shuttle conveyors to an overland supply conveyor. The overland conveyor would feed the secondary crusher, and the crushed product would be moved to any of four storage silos using conveyors. The in-pit crushing and conveying system is a unique fit for the power plants and provided many interesting challenges for the coal mine to overcome. System integration, maintenance windows, equipment availability and the like had a major role in engineering the system to a power plant complex having no emergency coal stock- pile and limited plant storage. In crediting the designers and operators, the in-pit conveying system at the Wyodak Mine has delivered coal to the power generators since 1994 with little or no downtime. ENGINEERING CONSIDERATIONS Yearly production for the mine is based on 3.5 MTPY which requires about 1,550 TPH on an 8 hour, one-shift- per-day, 5-days-per-week basis. Considering efficiencies for starting and stopping, crew transport, loader trams, and operational delays, a nominal conveyor feed rate of 2,150 TPH or 72% system utilization was used in engineering design application. This ton-per-hour basis is used to size and commission the remaining components of the in-pit system. Coal quality requirements also played a role in equipment selection. The mine reserve grades vertically and coal from Upper and Lower benches needed to be mixed in any given proportion. This required equipment to be filly variable and capable of producing all or part of the 2,150-ton-per-hour requirement. Bisecting the 80-foot- high coal face for quality required construction of a high lift module capable of reaching a single overland conveyor 40 to 50 feet above the pit floor. Designers built a high-lift, adjustable and portable module specific for this application Having two bench systems capable of each producing the required feed for the power plants required computerized control processes to control on-line production. RSLinx, RSLogix, and RSView were chosen as the communication and control software, and a Data Highway Plus serves as the communication backbone. Each of the two in-pit benches are controlled with a data center equipped with Allen Bradley 5-40 processors. This

Jan 1, 2004

By Ottmar Burghardt

Sintering of iron ore has undergone a dramatic development world-wide. Today, the annual world production amounts to approximately 600 million tons. At present, the sintering process in competition with pelletizing, seems to offer certain advantages inasmuch as it allows the use of solid fuels. Also, its charging behaviour in the blast furnace is, in general, considered as being better than that of pellets. This is true, in particular, for the charging behaviour and the gas permeability in relation to the grain size. Moreover, instead of charging limestone into the blast furnace, for instance, the basic components which are necessary for the burden materials can generally be supplied in a simpler and more cost-effective way by using basic sinter instead of basic pellets. While in sintering, with basicity rising to the range of 1.6 to 1.8 CaO : SiO2, the capacity increases and the sinter quality shows improvements in most of its characteristics, the pro¬duction of basic pellets generally requires a higher fuel rate, when compared to acid pellets, while production capacity decreases and some of the quality characteristics deteriorate.

Jan 1, 1982

By B. A. Ferguson, P. W. MacMillan

INTRODUCTION Jack Spalding in Deep Mining, Chapter 3, states: "In deep mining, to start stoping an orebody without a definite plan of operations covering the whole extraction from beginning to end is to invite serious trouble. Any stoping without plan is bound to leave, toward the end, a number of pillars, remnants, or promontories, which, as they are reduced in size, are liable to fail and cause a general collapse. The object of stope planning is so to mine an orebody that at no stage in the operation is a remnant left. Definition: `When a block of ore is stoped in such a way that eventually a small piece is left entirely surrounded by stoping, that piece is termed an island remnant.' The design of and adherence to a plan of stoping is of greater importance in preventing rock bursts than the type of ground control adopted." This, of course, is just an example of the old axiom that an ounce of prevention is worth a pound of cure. And in this context, mining does not have to be very deep before poor ground conditions, made worse by lack of good planning, force a mining company to resort to more expensive methods of ore recovery. At all times when rock pressures become excessive, stoping plans must be devised to avoid the creation of pillars or promontories. This is most readily recognized when mining deep narrow ore veins by open methods. It would be a little more general to say that stoping systems should seek to avoid or minimize localized concentrations of rock stress. Ore bodies, and particularly those at Falcon¬bridge, can reach large widths and leaving pillars for extraction at a later time cannot be avoided. However, good planning can reduce the effect of factors which create bad ground conditions. The necessary ingredients of planning are time and information: time to formulate and revise ideas; information, gained in exploration and early development, to provide the fullest knowledge of the ore body. In this way, plans become meaningful. It must always be kept in mind that the long-term results of good planning often require decisions not compatible with reaching earliest full production. PLANNING FOR DEEP MINING At Falconbridge Nickel Mines ore bodies presently being found and prepared for production lie at depths in excess of 610 m (2000 ft) below surface. Current un¬derground exploration is being directed towards favor¬able locations below 914.4 and 1219.2-m (3000 and 4000-ft) depths, at most mines. The Onaping Deep ore body has been traced to a depth of 1508.8 m (4950 ft) and at Falconbridge mine the ore body is under develop¬ment below the 5000 level. Planning for more efficient ground control has therefore become an important con¬sideration of future mining. HYDRAULIC BACK FILLING AND CUT-AND-FILL STOPING The advent of hydraulic back filling has led to the increasing use of flat-back cut-and-fill stoping methods during the last ten years, until it has probably become the commonest method of mining deep ore deposits in Canada. These methods account for almost all the pro¬duction from Falconbridge Mines at this time. The ad¬vantages claimed are: safety-a minimum of open ground; very little supplementary ground support (except rockbolted backs and walls); greater selectivity and flexibility in extraction (horizontal extension of the vein, high grade ore stringers going out into the walls); and high productivity and increased mechanization in larger ore widths. There is some difference of opinion as to whether hydraulic classified tailing back fill provides any great resistance to initial closure of the stope walls. The evi¬dence available suggests that it does not. Addition of portland cement to the back fill just prior to placement has produced a material having the properties of a weak concrete. It is confidently expected that this material will assist greatly in the efficient mining of pillars and generally in giving increased ground support to oppose closure. PRINCIPLES OF STOPE PLANNING BY LONGITUDINAL FLAT-BACK CUT-AND-FILL Again quoting from Spalding: "The practice of con¬trolling output by stopping and starting stopes is bad¬stopes once started should proceed without interruption, and control must therefore be obtained by altering the rate of stoping. It is therefore necessary that faces should normally advance at a medium rate which can be boosted, if required, to give temporarily a greater output. Modern high productivity stoping methods, however, generally aim to achieve and maintain the maximum output possible from a stope. Control of tonnage and grade, therefore, becomes a more complex problem requiring very careful scheduling and organiza¬tion. With the basic method of flat-back overhand stoping using hydraulic fill, it is not usually possible to mine in such a way as to avoid creating pillars." Consider an ore body as illustrated in Fig. 1. It is 274.3 m (900 ft) high by 228.6 m (750 ft) long on strike and averages 4.6 m (15 ft) in ore width. It rep¬resents approximately 907 000 t (1,000,000 st). The ore body could be mined by advancing one long single face through all the levels from bottom to top. A total face length of 228.6 m (750 ft) would accommo¬date five 45.7-m (150-ft) stopes. See Fig. la. Each of these stopes would produce approximately 1814 t/m (2000 stpm) and the ore body would require a period of ten years for complete extraction. A possible im¬provement to this scheme might be to stagger the indi¬vidual stope faces. See Fig. 1b. Problems of hanging wall failure can be reduced by this arrangement. From the point of view of reducing travelway maintenance, it is good practice to employ a stoping system of retreat. Stopes farthest from the shaft are mined first so that on completion, haulage drifts below can be aban

Jan 1, 1982

By N. Edmonson

During the last 30 years, a dramatic increase in the supply of byproduct sulfur substantially reduced the share of the market held by Frasch-mined (non-by-product) sulfur in North America. Even though Canadian sulfur production is not discretionary, the geographic remoteness from the large US markets and the high transportation costs associated with delivering sulfur to those markets has meant that Canadian producers act like discretionary producers in many important ways. They have effectively replaced the US Frasch producers as the cost-marginal sulfur source to large US markets. Implications of this change in market structure include more volatile sulfur prices and longer sulfur price recessions.

Jan 1, 1998

By Lawrence M. Brennan

Early in 1971 the Peabody Design and Preparation group began engineering on a preparation plant which was to process eight mil- lion tons per year of raw coal. For the first five or six years this product was to be furnished 50% from surface mining operations and roughly 50% from underground operations. By early 1980, 1008 of the production was to come from underground mines. Because of the large annual commitment from Randolph Preparation Plant, it was essential that the plant operate with minimum down-time. As usual, capital costs were to be kept at a minimum. Before we discuss the details of design, it would be best to discuss the overall picture. Underground production to be pro- cessed through Randolph Preparation Plant was to be delivered by means of one mile overland conveyor from Baldwin #1 Underground Mine to the plant. Surface Mine production was to be delivered by means of 100-ton haulage units dumping in a 1,000-ton capacity hopper. Because of the changing production rates in both the underground and surface mines as well as the great difference in size consist, it was decided that it would be necessary to divorce the plant as much as possible from production. For this reason, twin 6,500-ton raw coal silos were constructed to separate the two coals as well as allow blending. The preparation plant itself, was to be designed to Operate two shifts per day, five days per week, at an average rate Of 2,200 tons per hour with a maximum rate of 2600 to 2800 tons per hour. Coal was to be delivered to our customers by means of rail haulage at both unit train rates and plant process rates. The surface mined coal, after discharging into the 1,000-ton R.O.M. hopper, is fed by means of two 72" wide reciprocating feeders onto a 72" wide steel cable belt. The steel cabled conveyor discharges the R.O.M. coal onto a fixed grizzly, scalping out a percent of the 6 x 0 coal. The balance is fed to a 12' x 27' Rotary Breaker. The material passing through 6" round hole screens is added to the 6" x 0" from the scalping screen. The 6" x 0" is conveyed to the top of one of the 6500 ton silos designated for surface coal. Material that is not broken to pass the 6" breaker screen is discharged to a bin and disposed of with plant refuse.

Jan 1, 1976

By G. Brofas

The reforestation potential of bauxite mining spoils by seeding was examined on this research work. The experimental plot was established on the location Strofi Koukouvista 5, Ghiona Mountain. The experiment consists of 3 species (Pinus nigra, Robinia pseudoacacia, and Spartium junceum) with 4 sowing treatments and 2 sowing seasons (autumn and spring). The treatments were: (1) broadcasting; (2) broadcasting and soil cultivation; (3) broadcasting with straw mulching and straw binding with asphalt emulsion; (4) broadcasting with soil cultivation, straw mulching and straw binding with asphalt emulsion. Third year results show that spring is better than autumn as a sowing season. Only the Spartium junceum species were found to germinate and survive in autumn. The results are better for all the species on the spring seeding. The most effective treatments for Robinia pseudoacacia and Pinus nigra are 3 and 4 while for the Spartium junceum even simple broadcasting shows good results although the best ones are achieved with treatments 2, 3 and 4. The paper?s findings show that reforestation by sowing could be effective for some species but further experimentation with other species is recommended. Concerning Robinia pseudoacacia, testing of different seed preparation treatments is needed in order to maximize the germination of seeds.

Jan 1, 2002

By John Kuyt, Zili Li, Michael Mooney, Michele Mangione

"INTRODUCTIONTwin bored tunnels are commonly used for a variety of road and tunnel projects throughout the world. A major component of these twin tunnel projects is the construction of connecting cross passages that provide emergency access between tunnels and space for tunnel service equipment. Cross passages can present significant geotechnical risk and technical challenges to a project – support systems must be considered for redistributing loads around the mainline tunnel openings and supporting new loading resulting from excavation activities. Additionally, cross passages pose risks to construction schedule and cost because of their sequencing at the end of construction, so mitigation of schedule overruns can be critical. By observing field instrumentation data records from completed projects, a betterunderstanding of the force development in each temporary structure can be achieved with indication of which structures are the most critical to support ground loads.This paper discusses the data obtained during the Brisbane Airport Link construction project, a road tunnel located in Brisbane, Australia. Strain gauge instrumentation data from the precast segmental lining and propped opening steel sets is presented with relation to the construction sequence and local geology. Interpretations and conclusions based on the observed trends are proposed, with several key mechanisms affecting the loading and unloading identified.Brisbane Airport Link Project OverviewThe Brisbane Airport Link project in Brisbane, Australia, connects central Brisbane to the airport located northeast of the downtown with over 7km of twin-bore road tunnels, two lanes each, with several caverns, interchanges, and access ramps constructed along the alignment to connect to other key routes in the city. The project was executed under a Public-Private Partnership (P3) scheme, funded and operated by the consortium BrisConnections. The main contractors were a joint venture of John Holland and Theiss, while the design was carried out by Arup and Parsons Brinckerhoff.Construction of the tunnels was conducted using a combination of 17 roadheaders and two 12.5m double-shield Herrenknecht TBMs [1]. The TBMs were used in EPB-type closed mode for the initial portion of their drive, starting at the northeast end of the alignment. Tunneling started in soft ground, weathered rock, and mixed face conditions initially, progressing to competent sedimentary rocks as the drive continued. Support was accomplished using a 12.1 m diameter steel fiber reinforced concrete (SFRC) segmental lining installed behind the TBMs in a 9 segments + 1 key arrangement. The innovative design of these segments is detailed in the paper by Harding and Francis [2]. Ancillary roads and underground interchange caverns were constructed by sequential excavation techniques using roadheaders, accounting for nearly half of the total underground excavations."

Jan 1, 2016

By Ronald R. H. Cohen, Margaret W. Staub

Two pilot scale bioreactors designed to treat acid mine drainage primarily using sulfate reducing bacteria were constructed underground at the Eagle Mine Superfund site in Minturn Colorado. The project was designed to test th applicability of microbiological treatment on water with high metals concentration and to test the process in an underground environment. Objectives were, 1) to compare metal removal, pH increase, and sulfate reduction in a single stage reactor with a double stage reactor of equal volume at varying flow rates; 2) to determine where within the substrate treatment occurs· and 3) to examine volume versus surface ara as a design parameter for optimizing metal removal processes. Results from the reactors showed that the single stage reactor and double stage reactor performed equally well for the removal of Cd, Cu, Pb, Fe, Mn, and Zn. pH increased above 6.5 in both reactors. Greater than 97% of metals were removed within the first ten inches of substrate. Volume proved to be a more useful parameter for determining removal rates than area. Using volume as a design parameter rather than surface area enables PMDT systems to be constructed as packed columns rather than broad wetlands, Thus, space requirements are significantly reduced.

Jan 1, 1993

By Lj. Obradovic

The aim of this paper is to study the feasibility of copper and gold recovery from the old flotation tailing dam at Bor, Serbia. Laboratory leaching experiments were carried out using a composite sample taken by drilling from 16 - 40 m depth. Different oxidation agents as FeCl3, Fe2(SO4)3 and H2O2 were used. An H2SO4 solution with a pH less than 3 was used for leaching. The influence of a number of parameters on the recovery of both elements as temperature, pH, duration of leaching and concentration of oxidation agents was also studied.

Jan 1, 2006

By Robert E. Haun

Retech's Plasma Arc Centrifugal Treatment (PACT) system is a thermal technology that uses heat generated by a plasma torch to treat hazardous waste containing metals and/or organics. Metal bearing solids and soil are melted by the process, and organic contaminants are thermally destroyed. The molten material forms a hard, glass-like leach resistant mass after cooling in a mold. This paper discusses plasma processing in general and presents some results of treating mineral processing wastes using the plasma arc centrifugal treatment process.

Jan 1, 1995

By Harold B. Goldman

On the basis of tonnage, the sand and gravel industry is the second largest nonfuel mineral industry in the United States. In 1990, the production of sand and gravel was 927 Mt valued at $3.4 billion. California, which leads the nation with more than 126 Mt, together with Texas, Washington, Michigan, and Ohio, account for 36% of the total production in the nation (Table 1). In commercial usage, sand applies to rock or mineral fragments ranging in size from particles retained on a No. 200 Sieve (0.074 mm openings) to those passing a No. 4 Sieve (4.76 mm openings). Gravel consists of rock or mineral fragments larger than 4.76 mm, ranging up to 88.9 mm maximum size. The construction industry consumes 97% of the sand and gravel produced; the remainder is sand used for specialized products such as glass (see chapter on Industrial Sand and Sandstone and Glass Raw Materials). Utilization The building industry uses sand and gravel chiefly as aggregate in portland cement concrete, mortar, and plaster; the paving industry uses sand and gravel in both asphaltic mixtures and portland cement concrete. Aggregate is commonly designated as the inert fragmental material that is bound into a conglomerate mass by a cementing material such as portland cement, asphalt, or gypsum plaster. Sand and gravel is also used as construction fill, road base and subbase, and decorative material. Portland Cement Concrete Aggregates: Portland cement concrete consists of sand and gravel surrounded and held together by hardened portland cement paste. Concrete mixes commonly contain 15 to 20% water, 7 to 14% cement, and66 to 78% aggregate. Sand and gravel used as concrete aggregate have to meet many requirements (Goldman and Reining, 1983). Premature deterioration of concrete has been traced in many instances to the use of unsuitable aggregates. Asphaltic Aggregate: Asphaltic mixtures used predominantly for paving consist of combinations of sand, gravel, and mineral filler (material finer than 0.076 mm), uniformly coated and mixed with asphalt produced in the refining of petroleum. Sand and gravel used as asphaltic aggregate must meet the same general physical requirements as materials used for portland cement aggregate. GEOLOGY General Requirements of Aggregates Construction aggregate has many requirements that are difficult to meet if only unprocessed material from natural deposits is used. Suitable material is composed of clean, uncoated, properly shaped particles that are sound and durable. Soundness and durability are terms used to denote the ability of aggregates to retain a uniform physical and chemical state over a long period of time so as not to disintegrate when exposed to weathering and other destructive processes. Individual particles must be tough and firm, possessing the strength to resist physical stresses and chemical and physical changes, that may cause swelling, cracking, softening, and leaching. The aggregate should not be contaminated by excessive clayey material, silt, mica, organic matter, chemical salts, and surface coatings. Physical Properties: The quality of aggregate depends upon its physical and chemical properties. These, in turn, may be inherent mineralogical and textural features of the rock or may be the effects of later changes such as tectonics, mechanical or chemical weathering, or incrustations. The physical properties most significant for concrete use are: 1) abundance and nature of fractures and pores, 2) particle shape and surface texture, and 3) volume changes which may occur because of freezing and thawing or wetting and drying. An aggregate is considered to be physically sound if it is adequately strong and capable of resisting the agencies of weathering without disruption or decomposition. Minerals or rock particles that are physically weak, extremely water absorptive, and easily cleavable are susceptible to breakdown. The use of such materials in concrete reduces strength or leads to early deterioration by promoting weak bond between cement and aggregate, or by inducing cracking, spalling, or popouts. Severely weathered, soft, micaceous, or porous materials may cause localized stresses to develop in concrete by swelling and shrinking during wetting and drying or freezing and thawing cycles. Physical Suitability of the Various Rock Types. Sedimentary rocks have a wide range in physical and chemical qualities. Sand- stones and limestones, if hard and dense, are ordinarily satisfactory, but many sandstones are friable and excessively porous and commonly are clay-bearing. Shales generally make poor aggregate material, being soft, weak, and absorptive. Most igneous rocks are satisfactory, being normally hard, tough, and dense. Tuffs and certain flow rocks may be extremely porous and have high water absorption and low strength. Metamorphic rocks differ in character. Most quartzites are massive, tough, and dense. Fine-grained marbles are usually durable, but coarse-grained marbles have low abrasion resistance. Gneisses are ordinarily very tough and durable. Some schists contain micaceous minerals that are undesirable because they are soft, laminated, and absorptive. Micaceous minerals are susceptible to splitting along cleavage planes and thereby impair particle strength and durability. Some schists and slates in particular are thinly laminated and tend to assume flat slabby shapes that lack strength-and do not pack well. Any or all of these rock types may be rendered undesirable because of harmful exterior coatings. Weathering processes, particularly the action of ground waters, deposit these coatings. The most common coatings are calcium carbonate, clay, silt, opal, chalcedony, iron oxide, manganese oxide, and gypsum. Particles with these coatings are undesirable as aggregates because the bond between particle and coating may be weak, and decreasing the strength of the aggregate-cement bond.

Jan 1, 1994

By C. Pace, D. Darlington, A. Montera, C. Lingren

"A phosphate beneficiation plant in Utah is using isotope-based Prompt Gamma Neutron Activation Analysis (PGNAA) technology to analyze the entire slurry streams of the incoming feed and final product in order to adjust the flotation circuits in near real time and optimize the plant output. The plant has gained greatly improved visibility/control and has reduced product variability for improved profitability and product quality. This paper takes a look at the economics as well as application details and the difference this technology provides in a phosphate application. SIMPLOT PHOSPHATES INTRODUCTION Simplot’s Vernal, Utah Phosphate Mine and Mill are located in the northeast corner of the state in the Uinta Mountains along US Highway 191, eleven miles north of the town of Vernal. The phosphoric acid and fertilizer plant fed by the Vernal mine is located 94 miles Northeast of Simplot Phosphates in Rock Springs, WY.At the Vernal plant, loader and shovel mining to 85 Ton capacity trucks is done at an annual rate of 3.8 Million tons of ore. Run of mine ore is crushed in a mineral sizer to minus 14 inch diameter at the rate 11,500 tons a day and transported to the SAG Mill’s 130,000 ton surge stockpile by overland belt. ORE RESERVES The ore is a Phosphoria Formation of sedimentary rock type dating to the Permian Age. The ore body rests over a 25 mile area on Weber Sandstone from the Pennsylvania and Permian geologic period. Layers of Phosphate Rock can vary in thickness from 0.3 up to 2.1 feet, alternating with layers of Mudstone. The eastern edge of the ore’s body Panel “G” runs 16 feet deep with a feed grade of 21.5% P205 with lower MgO diluents of less than 2.79%."

Jan 1, 2016

By J. Lovering, H. Wilhelm, P. Siewert

Introduction The Afton property is located 290 km (180 miles) by air east-northeast from Vancouver and 14 km (8.7 miles) west of Kamloops, a city of 60,000 people, in south central British Columbia, Canada. The mine is adjacent to the Trans-Canada Highway at an elevation of 670 m (2198 ft) above sea level. The ore body is a porphyry copper deposit that has undergone supergene alteration. The major economic minerals in the supergene zone are native copper and chalcocite with chalcopyrite and bornite in the hypergene areas. The grade is 1% with an overall copper distribution - 70% native, 25% chalcocite, and 5% chalcopyrite with bornite and covellite. The ore also contains important but variable amounts of gold and silver. The mill was designed to treat 6350 t/d (7000 stpd). Semiautogenous grinding was selected to minimize capital cost and because of the expected high clay content of the ore, which would have caused problems in a conventional crushing and screening plant. Test work indicated that a recovery of 87% was possible in a circuit incorporating both flotation and gravity separation. Flowsheet Run-of-mine ore is crushed in a 1.06 x 1.65-m (3.5 x 5.4-ft) Allis Chalmers gyratory crusher set at 228.6 mm (9 in.), closed side setting. The surge pocket, below the crusher, is emptied by a Hydrastroke feeder onto number one conveyor, which discharges onto a 180,000-t (198,416-st) coarse ore stockpile. Six Hydrastroke feeders on two conveyors withdraw the crushed material from the bottom of the pile. These two conveyors, in turn, discharge onto the belt feeding the semiautogenous mill. The live storage in the stockpile is approximately 22,000 t (24,250 st), sufficient for three days' mill feed. Primary grinding is accomplished in an 8.5-m (28-ft) diam by 3.7-m (12-ft) long Koppers (Hardinge Cascade) mill (Fig. 1) containing a 10% ball charge and driven by a 4000-kW dc variable speed motor. The mill dis¬charge is pumped by a 10 x 12 G.I.W. pump to a 1.22 x 4.88-m (4 x 16-ft) stationary screen sloped at 20°. Screen oversize returns to the semiautogenous mill (SAM), and the undersize flows by gravity to the ball mill discharge pump box. Secondary grinding is performed in a 5-m (16.4-ft) diam by 8.84-m (29-ft) Koppers overflow ball mill driven by a 3430-kW synchronous motor through an air clutch. The mill is in closed circuit with a Krebs Cyclopac containing 10 635-mm (25-in.) cyclones and the cyclone overflow, at 35% solids and 65% to 70% -200 mesh, is flotation feed. In order to limit the buildup of native copper, circulating in the secondary grinding circuit, a portion of the underflow from the cyclones is processed in a circuit containing screens, cyclones, and shaking tables to produce a finished metallic copper concentrate. Primary mill variable speed drive The overall waste to ore ratio at Afton was 4.5:1. The mining was to be done with only three shovels, which meant that it was highly unlikely that more than one of them would be in ore at any one time. The resulting inability to blend the mill feed made it impossible to prevent wide swings in the grade and grindability. The variable speed do drive motor installed on the semiautogenous mill was selected because of the extreme variability of the Afton ore body. This variability has persisted throughout the lifetime of the mine. There are times, however, when due to ore conditions, the mill is operated at full speed (78% of critical) for extended periods of several shifts duration. There are other times when the mill speed may be changed several times in a 12-hour shift due to changing ore conditions. When ore is processed that contains a fairly large proportion of fine native copper, the primary mill speed and, consequently, the tonnage may be reduced to improve the secondary grind and to maintain an acceptable grind and recovery. High clay ores require less mill speed and more dilute grinding densities. In the latter case, the slower primary mill speed also helps to minimize damage to the mill liners. Approximately 57% of the time the mill operates between 90% and 100% of full speed or between 71% and 78% of critical. The variable speed is also used for inching during mill relines.

Jan 1, 1987

By Howard J. Haselhuhn

The effectiveness of selective flocculation and dispersion processes for the concentration of hematite ore are strongly dependent on the ionic content of the process water. It has been noticed that magnesium ions are more detrimental to iron recoveries during gravity separation and flotation processes than calcium. This phenomenon was studied by measuring the zeta potential of hematite ore at various concentrations of sodium, magnesium, calcium, strontium and barium at a pH of 11. Results show that zeta potential inverts from negative to positive at concentrations as low as 0.16 mM of magnesium ions in the solution. It takes a significantly higher concentration (1.25 mM) of calcium ions to achieve the same effect. This difference is attributed to the ability of magnesium ions to adsorb to all surface hydroxyl groups, whereas calcium, due to its larger size, can only adsorb to every third hydroxyl group. This hypothesis was confirmed by results similar to those seen with calcium when this test was repeated with strontium and barium ions. The tendency for calcium ions to adsorb to every third surface hydroxyl group causes higher concentrations of calcium to be less detrimental to hematite concentration processes than magnesium.

May 1, 2012

By R. M. Smailer

Detailed description of the development of the FIOR process1 and, subsequently, other technical papers relating to the operation and start up of the FIOR plant2,3 are available in the technical literature. The Fluidized Iron Ore Reduction (FIOR) process plant constructed in the Matanzas region of Venezuela represents the first commercial application of this technology. The FIOR process was developed by Exxon Company during the 1950's and 1960's; and the facility in Venezuela was designed and constructed by Davy McKee with the plant start up in 1976. While there is a minority government ownership in the company, FIOR de Venezuela operates as a private sector company with the majority of the ownership being private capital.

Jan 1, 1981

By K. R. Price, M. Gauna

Jim Walter Resources No. 4 mine produces coal through retreat longwall mining at 610 to 670 m (2000 to 2200 ft) of depth. A three entry, total yield pillar, longwall development experimental region was established for three adjacent longwall panels utilizing 6.1 m (20 ft) wide gate section pillars. Presently, the first longwall panel is completed with roadway deformational behavior monitored for a ground control assessment of the mining system. Data illustrate that the total yield pillar gate section design at No. 4 mine is presently comparable to gate section designs having large equal size chain pillars. Ground conditions will continue to be monitored during the mining of subsequent longwall panels in order to determine the ground control implications from multiple panel extraction.

Jan 1, 1986

By Kostas Fytas

Canada still remains today a strong resource nation with a huge geological potential. Canada ranks among the top five leading countries in the production of 15 major minerals. Acid mine drainage (AMD) is a serious environmental problem that preoccupies the Canadian Mining Industry. Considerable amounts of money are spent every year in an effort to prevent or reduce the acid mine drainage production. It is estimated that across Canada there are a few thousand abandoned mine sites as a result of the last fifty years of mining. Most of these sites are acid producing presenting a serious threat to the environment. It has been estimated that the cost of stabilizing and reclaiming these sites across Canada will be over $5 billion over the next twenty years. Current practices for decommissioning acid producing sites are water covers (flooding), dry covers, revegetation and treatment of the acid effluents. This paper presents the technical and legal issues of mine closure and reclamation in Canada. The methods used in the revegetation of acid producing mine sites is also presented This paper also reviews the financial guarantee requirements for mine closure by the provincial legislations in Canada. A case study of stabilizing and reclaiming a sulphide tailings area in the Sudbury region is presented. It illustrates the challenges facing the Canadian Government and the mining companies in restoring and revegetating the closed and abandoned mine sites across Canada.

Jan 1, 2002

By K. F. Unrug

The identification of the major subsidence areas in the Western Kentucky Coal Field has been made to distinguish the subsidence phenomena on the basis of geographical location Further, this identification process is based upon selected critical geological, topographical, and mining features in the coal fields. The proposed identification of the regions is derived from a comparison of the subsidence potential in a particular area. The most likely severe subsidence areas are compared to those with the probable low subsidence severity. The combination of certain geological and mining parameters involved allows for the derivation of a classification of the subsidence potential. It is recommended that further site specific studies should be performed on this area.

Jan 1, 1986

By M. Hancer

The sign of the surface charge of selected potassium soluble salt minerals namely KClO4 , K2SO4 , KNO3, and KCl was determined from nonequilibrium electrophoresis measurements and found to be negative for each of the salt. Dodecylamine flotation of these salts revealed that the concentration of amine required for flotation increases in the order KClO4< K2SO4< KNO3<KCI. Precipitation of the amine collector colloid was not required for flotation of the oxyanion salts and in this case it appears that adsorption of the dodecylamine cation is sufficient to create the hydrophobic surface state. The order for flotation is explained in terms of surface hydration of these respective potassium salts based on contact angle measurements and FT Raman spectroscopic characterization of water structure.

Jan 1, 1999

By P. H. Lu, M. J. DeMarco, J. R. Koehler

Beginning in late 1985 and continuing through 1987, US Bureau of Mines personnel investigated longwall chain pillar and entry design in two and three-entry gateroad systems at a deep underground coal mine in central Utah. To evaluate their respective stability characteristics, four chain pillars and two longwall panels within the two separate entry systems were instrumented with Bureau hydraulic borehole pressure cells to continuously monitor vertical and horizontal pillar and panel stresses through adjacent panel retreat. In addition, supplemental entry closure information was obtained from sites located in the vicinity of the instrumented pillars. The findings presented in this report demonstrate the practicality of evaluating mine pillar stability using in situ methods and support three significant conclusions regarding ground control aspects of the investigated two-entry system: (1) Total areal entry loading was considerably less for the two-entry system, as opposed to the three-entry system, (2) a smaller percentage of opened ground in the two-entry system contributed to a marked reduction in rooffalls, and (3) lower areal loading in the two-entry system suggests improved ground conditions when mining the underlying seam.

Jan 1, 1987