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By Anders Sellgren, K. Hansson, A. Jedborn

In the mining industry, conversion of fine particles into a slurry is often part of the normal processing of the ore. Therefore. from a system's point of view, slurry transportation would be considered as a natural alternative, all the way from the working face to the final processing. In underground mines, there is sometimes a considerable inflow of ground water that has to be pumped out. When the mine dewatering installations are integrated with a hydraulic hoisting system, the cost of power needed to pump out the ground water can be excluded from the cost of hoisting, compared with other modes of transporting the solids to the surface. The chief advantage of hydraulic hoisting is that the hoisting capacity can be increased at comparatively low total cost without necessitating the sinking of new shafts. Alternative examples include deepening of the existing mechanical hoisting system, truck, or conveyor belts transport in ramps to the present level. These often become very expensive. The economic effectiveness of hydraulic hoisting and hydraulic design considerations have been discussed by Kostuik (1965) and Sellgren (1985, 1986) for both small shallow mines and large deep underground mines. Objective and scope The objective of this study was to demonstrate the feasibiliy of hydraulic hoisting in hypothetical examples of mine deepening. Underground grinding to a maximum particle size of about 700 um (about 25 mesh) is considered. The pumping of ground ore products with positive displacement plunger or membrane type of pumps at working pressures of up to 15 MPa (2200 psi) is a thoroughly tested technique (Venton, 1986). Hydraulic hoisting system A hydraulic hoisting system is described schematically in Fig. 1. The ore is mined, handled, and crushed conventionally. It is then ground wet underground. The slurry is then temporarily stored in an agitated tank and pumped to an agitated receiving tank on the surface. The ore is then further processed in the concentrator. The capacity of a pipeline is very high. For example, an annual production of 500 Kt (550,000 st) of ore can be handled in a 100-mm-diam (4-in.-diam) pipeline. Systems layout - economic effectiveness Schematic mine layouts are presented that represent situations where hydraulic hoisting often is the most economical alternative. A situation where the ore is trucked to the level of the existing hoisting system is shown in Fig. 2 (following page). With hydraulic hoisting, only a shaft for ventilation, pipe- line, and personnel are needed. In a first phase (Fig. 2). a hydraulic hoisting installation with crushing and grinding units, storage facilities, and pumps has been located at a level of about 400 to 600 m (1310 to 1970 ft) below the surface. With further development of the mine, the hydraulic hoisting

Jan 1, 1990

By Roger S. Dewey

1.1. INTRODUCTION Despite indications of a general economic recovery in the U.S. during the mid-19801s, mine operators have not generally shared in the benefits of these reported better times. Budgets have not rebounded, borrowing is tougher, parent companies' pockets are not as deep as in previous years, and margins are not any better -- if there is any left over at all. Managing capital equipment budgets is always a challenge, and the last five years of stretched equipment lives make these decisions more tenuous. Delays and avoidance of replacements due to tough times inevitably makes those decisions increasingly difficult and usually more expensive The purposes of this paper are to present. a brief refresher outline of the fundamentals of equipment replacement analysis, to suggest a number of papers where analysis methods are explained in greater detail. and to recommend the use of a widespread microcomputer software capability (1-2-3) as the easiest path to a much-needed solution for a very difficult, and usually under-supported management decision problem 2. REPLACEMENT DECISION POLICY Investments in equipment replacements must compete for capital with all other investment alternatives. Capital for replacements of ten represents a greater commitment than do the new projects which a company might handle in a given year. But. sadly, formal policies for standardized analysis of replacements rarely exist, even in major companies. Yet this policy can be very significant in determining the competitiveness of a firm. Whereas many replacement decisions are based on cost minimization, economists and financial analysts agree that wealth maximization or return on investment is the superior criterion. Though the difference may seem slight, agreement on a firm's financial objective is vital. Replacement policy should be straight- forward. have a logical basis, and a proper level of sophistication. There are several reasons to develop a policy for replacement decision-making: [I] 1 - equipment's optimal economic life is nearly always shorter than its physical life; 2 - mine life often exceeds equipment life, so management must face these decisions routinly 3 - replacement capital costs, on an undiscounted basis, can run 25 to 75% of original capital 4 - technological changes may result in a newer machine or system which can operate more efficiently, more safely. or less expensively 5 - estimates should be made consistently in the fundamental areas of life expectancy, availability, utilization. price escalation, taxes 6 - long-term and short-term budgeting without decent data and analytical capability is nearly impossible, or at least ineffective Policy development assumptions which must be considered include: the planning horizon, cost patterns, interest rates, availability of capital, replacement vs. expansion, technology advances.

Jan 1, 1987

By Randolph E. Scheffel, David C. Fitch, David K. Hogan, Mark R. Welsh

In Oct. 1981, Ranchers Exploration and Development Corp. began producing silver bullion from the Escalante mine and mill complex, located in southwestern Utah. Current in-place reserves are estimated at 2.42 Mt (2.67 million st) of ore, containing about 936 t (33 million oz) of silver at an average grade of 428.5 g/t (12.5 oz/st). The vein-type deposit is being mined underground through a decline that is currently 864 m (2,835 ft) in length, but will ultimately be at least 1.5 km (5,100 ft) long. The mine is completely mechanized with rubber-tired equipment, and is producing 680 t/d (750 stpd) of ore, five days a week, or 177 kt/a (195,000 stpy). Additional capacity is available. The cyanide leach mill and associated refinery are presently producing 4.35 t/month (140,000 oz/month) of silver bullion-an an¬nual rate of about 52.25 t (1.68 million oz). With minor modifications, the mill can produce up to about 70 t/a (2.25 million oz/year), and expansion is currently underway. History The Escalante property is located 68 km (42 miles) west of Cedar City, UT, on the west side of the Escalante Valley, where the vein forms a gentle ridge in an area of low, rolling hills. For years, the vein has been the object of considerable exploration, but never became more than a sporadic producer, primarily due to low silver prices. Heber Holt of Enterprise, UT, staked the original four claims in 1896; they were patented in 1910. During the next 48 years, limited exploration and mining activity was conducted on the property. Systematic evaluation and development did not, however, begin until March 1958, when Samuel S. Arentz, a prominent mining engineer and consultant, formed several joint ventures with mining and oil companies that conducted drilling, drifting, mining, and water drawdown tests on the property. However, it remained uneconomic, principally because of low silver prices and costs for pumping large flows of underground water. In 1975, Ranchers obtained an option on the property and in 1979 began a drilling program. As of May 1982, the company had drilled 70 core holes, bringing the number of exploratory holes drilled on the property to 138. Test mining began in 1979 as silver prices increased, and full mine development began in 1980. Mill construction began in Sept. 1980, and on Aug 29, 1981, the first ore was fed to the mill.

Jan 9, 1982

By Joseph Fernandez, Luminita Calin

INTRODUCTION The Tunnel and Reservoir Plan (TARP), commonly known as the Deep Tunnel Project, is the daring and unique concept developed by the Metropolitan Water Reclamation District of Greater Chicago (MWRDGC) as a two-phase solution to the problem of waterway pollution and basement flooding due to sewer system backups created by the combined sewer over-flows (CSOs). Phase I, primarily for pollution control, includes the tunnels, shafts, connecting sewers and structures. These Phase I facilities intercept, during wet weather events, the CSOs within the MWRDGC TARP service area. The tunnels and shafts provide temporary storage of the CSOs for subsequent treatment at Water Reclamation Plants (WRP) before discharging into waterways. Ultimately, upon completion of Phase II, the tunnels, located from 45 m(150 ft) to 100 m (330 ft) below ground, will convey the CSOs to reservoirs for storage until WRP capacity is available. Phase II, which includes the reservoirs, pumping facilities, additional tunnels and shafts, has been planned to serve a dual purpose. The reservoirs provide storage of excess flows during wet weather events, those beyond the WRP capacity. Following the wet weather events, the stored flows can be conveyed, at controlled rates, to the WRPs for full treatment. These excess flows previously would have resulted in both overflows into the waterways and eventually backups in the local sewer systems. The storage capacity of the reservoirs also will promote the full utilization of the conveyance capacity of the tunnels and shafts, which in turn results in optimizing the conveyance capacity of the existing sewer systems. Therefore pollution control is further enhanced by allowing for capture of a greater volume of CSOs and greatly minimizing backups in the local sewer systems. The TARP project consists of four systems: O’Hare, Mainstream, Des Plaines, and Calumet, covering a total area of 971 km2 (375 mi2) and serving the City of Chicago and 51neighboring communities as shown in Figure1. With only two contracts left before completion of the Phase I tunnels, the 12.7km (7.9mi) Torrence Avenue leg, to be awarded in 1997, and the 12.4km (7.7mi) Little Calumet leg unfunded as of yet, both part of the Calumet System, the pollution control phase of the project is now 86% complete (Table 1).

Jan 1, 1997

By Leonard G. Austin

INTRODUCTION Run-of-mine (ROM) coal generally falls into two major groups: that from underground mining (continuous mining machines) and that from strip mining. Continuous miners produce a finer product and strip mining a coarser product that is crushed to produce the desired size. Further size reduction depends on the proposed use or uses for the coal. For example, for pulverized coal power plants as a rough general rule, the coal is ground or pulverized to 80% less than 200 mesh (0.075 mm) before it is blown into the boiler furnace as a coal-air suspension. For solvent-refined coal processes, the desired size is about <30 mesh (< 0.60 mm); in fluidized bed uses, the coal is typically < 8, in. (< 1.6 mm); for Lurgi pressure gasifiers, the desired size is 1 to 2% in. (25 to 63 mm). Coal preparation plants typically work with a coarse size fraction of 6X 1/14 in. (150X32 mm). The past decades has shown a steady increase in the tonnage of coal cleaned, 1 but it seems probable that the next decades will produce much wider applica- tion of deep coal cleaning before use and that the stages of crushing and pulverization from ROM coal to end use will be integrated with coal cleaning processes. Since different cleaning processes work better with different ranges of coal size and the degree of cleaning possible is intimately connected with the size of coal, a systems design engineer will need as much information as possible concerning the size distributions produced by different types of machines and how these can be modified by different ways of operating the machines. In general, the following questions will be of concern: 1.What type of crusher or grinder is best, or most satisfactory, for reducing a given feed size and type of coal to a desired product size? Frequently, the desired product is smaller than some nominal size, but with as few fines as possible, or the fraction in some size range has to be as high as possible. 2. How big does the machine have to be for a given throughput rate? 3. How much electrical energy (or its equivalent) is required per ton of product? 4. How does the size distribution vary with change in throughput rate, and is a control scheme necessary or desirable? 5. Can the size distribution be readily varied if desired? 6. What is the optimum way to operate a machine in a given system? This section will emphasize certain general principles which guide us in answering these questions. PARTICLE SIZING Since particle size and size distribution are the raw information with which we work, it is worthwhile to review the terminology used. In this regard, see

Jan 1, 1991

By Andreas G. Mueller

Gold-silver and copper-gold deposits with high-temperature wallrock alteration of Ca-Mg-Fe-Al silicates occur in ca. 3.0 Ga old greenstone belts of the Yilgarn Block, Western Australia. The deposits are hosted by amphibolite-facies komatiites, basalts and BIF, controlled by ductile shear zones and located within the contact-metamorphic aureoles of composite granodiorite-granite batholiths. Disseminated sulfides and free gold form foliation-parallel ore zones in a gangue of anhydrous silicates (garnet, clinopyroxene, olivine), hydrous silicates (amphibole, biotite) and calcite. The deposits may be classified on the basis of their dominant economic metals as: 1(1) Gold skarns with substantial silver as by-product, represented by the Marvel Loch mine (production: 2,200,000 t at 3.5 g/t Au), and characterized by the mineral association pyrrhotite, loellingite, arsenopyrite, galena, gold and magnetite. (2) Gold skarns with low silver contents, represented by the Nevoria mine (production:1,850,000 t at 3.4 g/t Au), and characterized by the mineral association pyrrhotite, chalcopyrite, arsenopyrite, gold, Bi-tellurides and magnetite. (3) Copper-gold skarns, represented by the Elverdton mine (production: 850,000 t at .7 5 Cu, 0.81 g/t Au), and characterized by the mineral association magnetite, chalcopyrite, pyrrhotite, pyrite, gold and Ag-Bi tellurides.

Jan 1, 1990

By R. B. Steele

The paper contains the results of the physical and mathematical simulations (based on the finite elements simulation) of the length and shape influence on the mechanical strength of the brex. Key words: Stiff extrusion, extruded briquettes, BREX, mechanical strength, cold strength, compressive strength, tensile splitting strength, oval shape, finite elements simulation. INTRODUCTION Extruded briquettes (BREX) exhibit a set of unique properties when compared with briquettes formed by alternative methods. Brex generally have sufficient integrity right after extrusion to withstand the necessary handling and stockpiling and does not require any special means and tools for the green briquettes strength conservation. In a typical die without an ancillary cutting device green brex tend to have very high length to diameter (L/D) or length to section area ratios. For round hole dies the brex are typically 5to 6 times the diameters and even longer. Stiff-extrusion allows also for the very easy changing of the forming dies. Dies with different cross-sectional shapes of the brex (round, ovals, square, hollow structures and many other two dimensional shapes) are possible, and common. The results of our investigations show how these geometrical shape factors contribute to overall quality of the brex both in terms of strength and as new charge components in metallurgical furnaces and processes [2-4]

Feb 27, 2013

By C. P. Lazzara, R. W. Pro, A. C. Smith

The U.S. Bureau of Mines conducted a study to evaluate the effectiveness of automatic water sprinkler systems in the extinguishment of incipient conveyor belt fires under ventilated conditions. Large-scale experiments were conducted using a double strand conveyor belt configuration. Standard response, pendent-type sprinklers, with activation temperatures of 100°C, were installed above and between the two strands of belting, in accordance with Federal standards for sprinkler system installations in belt drive areas. Experiments at airflows of 1.1 and 4.6 m/s showed that the sprinklers activated later, the peak heat release rate was larger, and more betting was consumed at the higher airflow. In experiments with 74°C, fast response, directional sprinklers, the sprinklers activated at the same heat release rate for both high and low airflows, but the peak heat release rate and amount of bett consumed was slightly higher at the lower airflow. Experiments were also conducted with 100 °C, standard pendent sprinklers installed above the top belt, in accordance with the National Fire Protection Association (NFPA) Standard 123 for sprinkler system installations in conveyor belt drive areas in underground coal mines. In these experiments, the heat release rate and amount of betting consumed was larger at the higher airflow.

Jan 1, 1995

By B. H. Gardner, S. Serata, D. Dickie

Since 1958, the Sifto Salt Division of Domtar, Inc., has operated the Goderich (Ontario) salt mine in Michigan basin bedded halite, with shafts located on Lake Huron&apos;s eastern bank. Initial mining utilized a conventional room and pi1lar pattern. However, roof deterioration developed. Incidence of progressive upward failure of saltback and over1yings ha1ey-do1omite layers intensified with entry aging. Studies were therefore initiated on new mining patterns utilizing the Serata Stress Control Method. Pursuant thereto, three test areas, indicated in the mine plan of Fig. 1, were designed and excavated. Full conversion to Stress Control mining was effected in January, 1983. This paper focusses on the instrumentation work in the third test area. Test Area 3 was designed to validate the excavation geometry of the new stress-controlled main entry system. The testing of that design was accomplished, in part, by comprehensive in situ stress, property, and deformation measurements on the performance of the model main entry section of Test Area 3. The paper describes: (1) the theoretical background of the SPDR/Stress Control Method; (2) specific problems and objectives addressed by the 1983 program of in situ measurements, (3) measurement results; and (4) conclusions drawn therefrom.

Jan 1, 1986

By J. M. Dean, R. L. Grayson, N. P. Reddy, Y. J. Wang

This 1988 survey provided a cross section of computer software use by the US coal industry. From 1250 mailed questionnaires, 171 responses were received, representing a 4.9% sampling of the 3496 US coal mines. Since the distribution of responses by regional stratum matched well with the distribution of mines by stratum, the survey was given statistical validity. A statistical description details the computer use by region, mining type, organizational size, the level of application within the organization, the type of software application used, the variety of computers on which the software was used and the percent of users within engineering, management and general applications categories. Overall, 63.5% of the respondents use computers at some level in their organizations. All of the mines producing more than 454 kt/a (500,000 stpy) used computers. However, for those producing more than 9 kt/a (10,000 stpy) but less than 91 kt/a (100,000stpy), only 44% used computers. Generally, the software applications were still data-processing oriented. The application sophistication grew with increasing mine size. Computer system models and configurations for different sized organizations are recommended based on the configurations used in the industry at the time of the survey.

Jan 1, 1994

By Philippe Dumont

Presentation of the project Underground civil works to be performed They include two single track running tunnels and a service tunnel of a length of fifty kilometers . Many cross adits (180 adits) : to link the tunnels (53 cross passages), to relieve air super pressure from one tunnel to the other (87 piston ducts), to install technical adits (40 electrical rooms). Two cross-overs located at one third of the whole length, allow trains to shunt from one tunnel to the other so as to ensure one way traffic in case of emergency. Pumping stations allow water to be collected at the lowest point and pumped out. 1 Main features of the tunnels 1.1 The service tunnel This 4.8 m inner diameter gallery allows During civil works - front probing operations of the geological layers encountered (front and vertical probes). - lateral soil grouting that may be necessary to facilitate the boring of the two running tunnels . During operations - Maintenance and safety vehicles traffic - Fresh air supply to running tunnels - Fire protection and drainage system - Electricity supply and control and communication. The service tunnel has been bored with a 5.72 meter external diameter Robbins TBM for the undersea section, and a MITSUBISHI TBM for the land section.

Jan 1, 1991

By Nigel Pearson, Malcolm J. McPherson

In the Grasberg Mine, a series of orepasses of diameter 2 to 4m are being constructed, each one to carry as much as 10,000 tlhour of copper- and gold-bearing ore through single stage vertical drops of over 500m. At the base of each orepass, the fragmented rock falls into an orebin of diameter 1 Om and 100m deep. Long before the planned tonnages were reached on the early orepasses, a problem arose that was both serious and intriguing. Pressurization of air in the orebins caused by the falling rock reached such levels that when the feed to the top of an orpass was interrupted, a massive surge of air ascended rapidly and produced a dangerous backblast accompanied by dust; grit and rock fragments from the top of the orepass. Backblasts lasted for about one minute and were not only hazardous to the health and safety of personnel but also required time-consuming clean-up operations. The paper describes the aerodynamic analyses that were con- ducted in order to understand fully the mechanisms involved in the phenomenum, and the development of a procedure that al- lows the effects of falling material to be incorporated into existing ventilation network software.

Jan 1, 1997

By C. J. Pritchard, A. B. Cecala

An evaluation was performed on a district recirculation system at a trona mine in Granger, WY. A 1.52 m-diam (5 ft-diam), 37.3-kW (50-hp) vane axial fan was installed in a crosscut between the intake and return airways and recirculated about 31.4 m3/s (66,578 cfm) of air back into the intake. This amounted to a recirculation factor of about 20% of the return air. This air was part of the primary ventilation to three different continuous miner sections at operating distances of between 610 m (2000 ft) and 1829,m (6000 ft) from the recirculation fan. The mine environment was constantly monitored for respirable dust, methane and carbon monoxide in the operating sections and in the main intake and return entries. In addition, air samples were taken with and without the recirculation system during the evaluation and analyzed using gas chromatography. The results of this study showed only small increases in dust, while all other mine contaminants remained at such low levels that it was impossible to determine any relative changes from the recirculation system. Because of this, a brief SF6 tracer gas analysis study was subsequently performed to more precisely determine an increase in a mine contaminant with the recirculation system.

Jan 1, 1992

By Bernard F. Anderson

INTRODUCTION In this chapter, the term "down-the-hole drill" (DTH drill) is used as a generic name that encompasses the various trade names and other references such as "downhole drill," "in-the-hole drill," etc. This chapter is limited to a description of DTH drills used in stoping large underground ore bodies. DTH drills differ from conventional drills by virtue of the placement of the drill in the drill string. The DTH drill follows immediately behind the bit into the hole, rather than remaining on the feed as with ordinary drifters. Thus, no energy is dissipated through the steel or couplings, and the penetration rate is nearly constant, regardless of the depth of the hole. Since the drill must operate on compressed air and tolerates only small amounts of water, cuttings are flushed either by air with water-mist injection or by standard mine air with a dust collector at the collar. HISTORICAL DEVELOPMENT Mine managers have long known the economies enjoyed by quarry and open-pit operators in producing large quantities of ore. The savings are due primarily to the availability of massive equipment, capable of drilling large blastholes to reduce the amount of drilling, increase the fragmentation, reduce secondary blasting, and im¬prove the flow of the product. In an attempt to reduce underground mining costs, various methods are used for long-hole drilling, includ¬ing standard pneumatic percussion drifters and diamond drills. These systems have their shortcomings; percus¬sion drills are limited to small hole sizes and they ex¬perience excessive deviation and significant loss of energy with increased depth. The diamond drills provide deeper and straighter holes, but only at high cost. Both systems suffer from high noise levels, low penetration rates, and poor explosives distribution, among other problems. When the mining companies approached the drill manufacturers for a compact and portable large-hole jumbo for underground use, they specified not more than 1 % deviation on 60 m (200 ft) of vertical hole and a penetration rate of 15 m/h (50 fph). On Dec. 23, 1960, a test unit was placed in service in Montana and met the performance criteria. Though lacking the so¬phisticated features available today, the economies of surface blasting were brought underground. Unfortunately, the first system did not gain immedi¬ate acceptance in the industry. Among the factors con¬tributing to its demise were resistance to change, the need to alter development methods for the ore bodies, and a lack of flexibility in moving the rig from setup to setup and from level to level. In 1972, the mining industry again challenged the drill manufacturers to provide a workable jumbo that would combine compactness, ease of maintenance, relia¬bility, and efficiency, all on a self-propelled chassis. The manufacturers responded by providing improved jumbos, which have been accepted with enthusiasm throughout the mining industry. Today&apos;s DTH jumbos are capable of drilling from 100 to 200 mm (4 to 8 in.) diam holes that can be reamed to even larger diameters. The holes can be drilled to depths of 150 m (500 ft), depending upon ground conditions and the capability of the jumbo to retrieve the steel and drill. Fig. 1 illustrates a typical DTH jumbo. APPLICATIONS The uses to which DTH drill jumbos have been put are quite numerous, with new uses being found regularly. For convenience, these uses may be classified as primary blastholes and nonblasting holes. Primary Blastholes The original purpose for the development of the DTH jumbo was for drilling primary blastholes that could be mined by open-stope methods. Prior to the advent of the DTH jumbo, extensive development was required before production drilling could begin. Sub¬levels were required to allow access for column-and-arm stopers or ring/fan jumbos, to the extent necessary based on the effective penetration of the chosen machine. With the DTH jumbo, the mine engineer is able to reduce preproduction time and development costs. How¬ever, the most significant saving results from an im¬proved cost per ton of broken ore in the production phase. To utilize a DTH system, only a top heading and drawpoints are necessary. The top heading can be the width of the ore body with a 3.7-m (12-ft) back. A drop-raise pattern is drilled and shot to begin the stoping operation, providing a free face for subsequent blasting. A typical layout is illustrated in Fig. 2. The advantages of this system include: 1) Drilling and blasting are independent operations, and blasting can be performed at a rate congruous with the mine&apos;s ton-per-day capacity. 2) The development layout is simplified. 3) Good explosive distribution is achieved, provid¬ing more uniform fragmentation. 4) Environmental conditions for operators are im¬proved, including improved safety with all work directed downward (not overhead), lower noise levels, little fog, and a reduced dust count. 5) Improved production per manshift. 6) Simplified and easier operator work cycles. 7) Reduced cost per ton of product. 8) Fewer holes lost due to ground shifts. Nonblasting Holes With the introduction of the compact DTH jumbos, other practical uses became apparent, including the drilling of: 1) Holes for sand fill, from level to level and from level to stope. 2) Drain and dewatering holes. 3) Power and communications cable holes.

Jan 1, 1982

By Donovan F. Symonds

INTRODUCTION The selection of the optimum coal preparation plant flowsheet is a vital part of the development of a new plant or the expansion of an existing facility. Selection of an inappropriate flowsheet or individual equipment items can result in costly production losses, or needless expenditures or redundant or oversized equipment. The accuracy of the flowsheet selection process is, like most decisions, heavily dependent upon the amount, reliability, and quality of the database. A flowsheet which is based on simple borehole information from an area that has experienced little or no previous mining activity must necessarily include significant allowances for uncertainties. However, a plant which is to be built at an existing mine with good information on run-of-mine (ROM) coal quality, as well as bulk washability data and accurate projections of future quality, can be optimized. The following chapter will describe the processes which should be followed to select the most appropriate flowsheet for each unique operation. They include: 1. Examining and adjusting the washability database to reflect the ROM feed. 2. An evaluation of clean coal specifications. 3. Selection of the level of washing required. This includes no washing, if the coal is already at the desired specification, all the way through to complex total washing circuits for metallurgical grade coals. 4. Selection of the appropriate circuit such as dense medium bath vs. Baum jig. The most appropriate circuit is generally determined on the basis of economics. An approach for performing the evaluation will be desired. 5. The steps that need to be followed to develop a balanced plant flowsheet. ADJUSTING WASHABILITY DATA It is not unusual to have as many as 10 to 20 sets of washability information available for a particular mining property. These may be a combination of core and bulk samples or all bulk samples. In the latter case, they may also be face and/or ROM samples. It is necessary to reduce all of this information to a single set of washability data that is representative of the mean feed to the proposed preparation plant. This set is used to compare the clean coal yield and quality characteristics of the various flowsheet options. The term "representative" has some serious negative connotations. It is recognized that ROM coal quality will vary throughout a mine. However, over a period of 250 to 350 days per year, certain trends can be established. If possible, this "representative" sample should be close to the coal quality of the mean daily sample in terms of raw ash and sulfur contents, clean coal yield, and clean coal ash and sulfur content. If possible, this sample should be within the

Jan 1, 1991

By Shawna Von Stockhausen, Hong Yang, Tung T. Vu, Derek Penrice, Kevin Huynh

"The Regional Connector Transit Corridor is a 1.9 mile long light rail transit project in downtown Los Angeles, California. The project includes approximately 4,930 feet of twin bored tunnels together with a mined crossover cavern, 3 deep stations, and other cut-and-cover structures. The bored tunnels will be excavated using an earth pressure balance tunnel boring machine through Fernando Formation (extremely weak to weak rock) and alluvium. The tunneling will be performed in close proximity to a large number of existing structures, including the existing Metro Red Line tunnels with as little as 6 feet of separation. Thus, the evaluations of the tunneling effect on these structures and the instrumentation and monitoring program is an essential part of the project design. This paper provides a very brief overview of the considerations and evaluations of the tunneling effects on the existing structures and results of evaluation obtained during the final design stage of the project. INTRODUCTION The proposed Regional Connector Transit Corridor (RCTC) project provides a 1.9 mile-long connection between the Metro Gold Line from Pasadena and East Los Angeles; and the Metro Blue Line to Long Beach and the Expo Line to Culver City. The alignment is shown in Figure 1. The Regional Connector will provide new tracks through downtown Los Angeles. The project will extend underground from the Metro Blue, Expo, Red, and Purple Lines’ 7th Street/Metro Center Station following Flower Street, curving northeast to the north of 3rd Street under the 2nd Street Tunnel. The alignment will then continue eastward under 2nd Street, turning northeast to cross under the intersection of 1st and Alameda Streets, then splitting into two branches above the ground surface, one branch extending north along Alameda Street and the other extending east along 1st Street. The project elements include 3 deep stations, a SEM cavern, segments of TBM bored tunnels, and cut-and-cover guideway structures. The bored tunnels have an excavated diameter of 20.3 feet. The subsurface conditions along the tunnel alignment generally consist of artificial fill overlaying finer-grained alluvium (dominantly medium dense to dense sandy soils), underlain by coarser-grained alluvium (a wide variety of medium dense to dense sandy and gravely soils). Below the coarser-grained alluvium is Fernando Formation, which is characterized as extremely weak to very weak sedimentary bedrock. A long stretch of the alignment will be excavated in this formation. The soil layers above the Fernando Formation are of variable thickness and spatial distribution. The geologic profile is shown in Figure 2."

Jan 1, 2016

By William S. Pintz, William A. Vogely, Charles J. Johnson

This chapter on the role of minerals in developing economies is oriented toward economic policy issues. The discussion and analysis cut across the technical, economic, legal, and policy fields. The goal of this chapter is to pro- vide relevant background information important to individuals involved with the minerals sector in developing countries. Where possible, trends are shown to provide a basis for understanding the evolving role of minerals in developing economies. This chapter is divided into the following five sections: (1) characteristics of developing countries (which presents selected statistical comparisons); (2) government objectives in developing countries; (3) government policy formulation in developing countries; (4) determinants of future trends, and (5) a summary of the chapter&apos;s major points combined with the authors&apos; view. CHARACTERISTICS OF DEVELOPING COUNTRIES There are 155 nations in the world categorized as "developing countries." In total these nations represent about one third of the value of the world&apos;s mineral production, three quarters of the world&apos;s population, and over half of the world&apos;s land surface area. The group of so-called developing countries is quite diverse and there are exceptions to almost any general observation about the group as a whole. Most definitions of the term "developing country" are based on levels of per capita income. However, if per capita income is the sole criterion used in categorization, then several oil-exporting countries would jump from the category of developing country to the category of developed country.a Obviously, this results in a poor fit with respect to other measures. In general, the majority of developing countries can be characterized by relatively low per capita income, high population growth rates, high levels of unemployment, low levels of education, short average life ex- pectancies, income distribution skewed toward the highest 20% of the population, and a large share of exports represented by primary commodities. b For the purposes of this chapter a slightly modified version of the widely used World Bank system of classification (shown in Table 1.2.1) is used. As summarized in Table 1.2.1, developing countries are divided into three main groups based on estimated gross national product (GNP) per capita. It should be noted that the high- income, oil-exporting countries have characteristics of both developing and industrialized countries; however they are included for most statistical comparisons within the group of developing countries. Of the 155 developing countries in the world, approximately half produce minerals other than common building materials, and only 32 account for more than five percent of world pro-

Jan 1, 1985

By Tim Neil, O&apos

Finnish mining traditions date back 400 years. The country was then one of Europe&apos;s major iron producers. Most of the mines in current operation, however, were started during the last 20 to 30 years. The Finnish mining industry must make do with limited resources. Deposits are usually low grade and small in size. Open-pit methods are used where possible. But thin ore bodies and harsh winters can impose severe restrictions. Underground mining is often the only option. Despite these handicaps, Finnish metal miners produce copper, nickel, zinc, vanadium, iron ore, cobalt, and ferrochrome. Ironically, the success of Finnish mining is rooted in a crisis the industry faced several decades ago. Mining costs were rising. High grade mineral deposits were nearly exhausted. It was vital that Finland develop new mining machinery and methods that afforded higher productivity. Cooperation between the Finnish mining industry and its equipment manufacturers was good. Partly as a result, the Finnish mining industry has turned itself into one of the most modern and efficient in the world. The Finns did this while continuing to meet the stringent safety and environmental requirements in force in the Nordic countries. Today, the share of domestic raw materials used by the Finnish metallurgical industry is decreasing. So the country increasingly depends on imports of ores and concentrates. To better provide these imports, Finnish mining companies have expanded their foreign operations. And, while importing raw materials, the mining industry has been exporting its technological and equipment know-how and expertise. Mining methods Conventional mining methods and machinery are used in surface mines. Drilling is the exception. It is done with Tamrock hydraulic track drills and 100 to 150 mm-diam (4 to 6-in.-diam) drill holes. Pumpable slurry explosives are used in larger open pits. Underground, sublevel stoping with open stopes is the most common mining method. It accounts for about 80% of all underground stoping. The main benefits of this method are low production costs, high mechanization, and easy control. The best productivity figures exceed 10 t (11 st) per man-hour, including all mining personnel. The development of cable bolting has made it possible to use sublevel open stoping even in relatively weak rock. Cut-and-fill methods, such as concrete-pillar stoping, inclined-wall stoping, and traditional cut-and-fill have also been used, combined with sublevel caving and room-and-pillar methods, according to local demands. Various fill materials and stabilization components have been investigated and tested for cut-and-fill methods. Sublevel open stopes are often later filled, to stabilize the surface or to allow pillar recovery.

Jan 10, 1988

By J. B. Gwiazda

Introduction Thick seams constitute a considerable portion of Polish coal reserves. In producing mines and those under construction, the reserves where the height is more than 4.5 m (15 ft) total about 24v0 of all the hard coal reserves. The reserves that are under urban and industrial areas can be extracted using stowing methods for surface protection. About 50v0 of thick seam reserves can be extracted by caving methods. Thick seam mining methods Thick seams are usually extracted by multilift methods. There are some methods of thick seam mining that generally can be divided into caving and stowing methods. The stowing method is more expensive, but must be used when mining to protect surface features. The more known methods of thick seam mining include: • extraction of thick seam in ascending order with stowing; • thick seam mining in descending order with consolidated stowing; • thick seam mining with gob consolidation; • thick seam mining in descending order with roof caving; • sublevel caving thick seam mining system; and • thick seam mining system using an artificial roof. • Thick seam mining with introduction of an artificial roof The most advantageous method of thick seam mining is the method using an artificial roof. The artificial roof usually consists of wire mesh. Such wire mesh must be durable and must prevent small rock fragments from passing through. The mesh is usually made of steel properly protected against corrosion. The artificial roof is placed on the coal floor of the top slice (Fig. 1). Caving rock falls on the artificial roof and is thus separated from the coal. The next slice is extracted under the artificial roof. The producing face of the next slice must progress within 10 m (33 ft) behind the front of the first slice. As the second producing face progresses, the wire mesh moves down on the coal floor and may be used as an artificial roof for the next slice. Thus, one artificial roof allows the extraction of several consecutive slices. Application of the artificial roof allows for clean extraction and reduces coal losses. It also eliminates the hazard of spontaneous combustion. Introduction of the artificial roof has been limited due to a lack of artificial roof laying equipment. Manual laying of wire mesh is a labor-intensive operation, and it delays face progress, resulting in poor productivity. To eliminate these problems, artificial roof laying equipment was designed at the Polish Central Mining Institute. The support and the method of laying operation is covered by a patent. Artificial roof laying powered support To extract thick seams using an artificial roof; four-leg shield supports with lemniscate linkage were developed (Fig. 2). This shield is used in the first top

Jan 1, 1988

By M. R. Yalamanchili, J. D. Miller

Discussion by M.R. Yalamanchili and J.D. Miller The authors, M. S. Celik et al., should be recognized for their efforts to describe the flotation behavior of boron minerals. In the case of borax and other soluble salt minerals, analysis of the flotation chemistry has been difficult because of the high ionic strengths associated with these soluble salt systems. However, considerable progress has been made in this area, and recently a surface charge/collector colloid adsorption model was proposed by Miller and his coworkers to explain the collector adsorption phenomena observed in soluble salt flotation systems (Milleret al, 1992; Yalamanchili et al., 1993; Miller and Yalamanchili, 1994; Yalamanchili and Miller, 1994a: Yalamanchili and Miller, 1994b). In this work, the sign of the surface charge of alkali halides in their saturated brines was established on the basis of nonequilibrium electrophoretic mobility measurements by laser-Doppler electrophoresis (Miller et al., 1992). Generally, these results are what would be expected from the simplified lattice-ionhydration theory. This electrokinetic information coupled with the stability and prevalence of collector colloids in such soluble salt flotation systems indicates that the selective flotation of alkali halides is due to the adsorption of oppositely charged collector colloids by heterocoagulation. Experimental flotation/bubble attachment results for 21 different alkali halides (Yalamanchili et al., 1993; Yalamanchili and Miller, I994b) confirmed that the flotation response of soluble salt minerals with weak electrolyte collectors can best be explained by the adsorption of oppositely charged collector colloids rather than by the adsorption collector ions and/or neutral molecular dipoles as originally suggested by many researchers (Fuerstenau and Fuerstenau, 1957; Schubert, 1967; Roman et al., 1968). In addition, the flotation of certain alkali oxyanions (Pizarro et al., 1993) and double salts such as schoenite and kainite can be explained by the same collector colloid adsorption mechanism (Miller and Yalamanchili, 1994). The borax flotation results reported by Celik et al. need to be examined in terms of the above mentioned surface charge/ collector colloid adsorption model. Unfortunately, the authors seem to be unaware of this recent work that nicely describes soluble salt flotation with weak electrolyte type collectors such as amines and carboxylates. In view of our past work, the flotation characteristics of borax were of particular interest, and, in this regard, the results of dodecyl amine flotation of borax reported by Celik et al. have been examined in further detail in the light of experimental results from our laboratory. In our research, a vacuum flotation technique was used to study the flotation response of borax (Na2B407.10H20), which has a solubility of 39 g/L at 25 °C) with dodecyl amine hydrochloride as collector. These chemicals were purchased from Eastman Kodak and used as received. Saturated solutions of borax at desired pH values were prepared by continuously stirring the salt solutions over a period of about 10 hrs. It should be mentioned that the conditioning time to achieve equilibrium is an important variable and can significantly change the flotation response of some soluble salts (Yalamanchili et al., 1993). Collector was added to the saturated borax solutions containing about one gram of 100x 150 mesh borax particles, and conditioning was done for about 20 minutes prior to flotation. The borax flotation recoveries from saturated brine are presented in Fig. 1 as a function of collector addition at the natural pH of 9.3, as reported both by Celik et al. and as measured in our laboratory. In addition, the region of precipitation for the dodecyl amine hydroborate is included in Fig. 1. It can be seen in Fig. 1 that the flotation response curves are separated by about one order of magnitude in R12NH3CI collector addition. The flotation results of Celik et al. show that the maximum borax recoveries can be obtained below the solubility limit of the dodecyl amine hydroborate collector. However, in our experiments borax flotation seems to occur only after the precipitation of the dodecyl amine hydroborate collector as might be expected from the collector colloid adsorption model (Yalamanchili et al., 1993) if borax were negatively charged. Further analysis by nonequilibrium and equilibrium electrophoretic mobility measurements for borax indicates that borax is negatively charged at the natural pH of 9.3, as discussed below. The reliability of the nonequilibrium electrophoretic measurements has been demonstrated previously for alkali halides and alkali oxyanions (Miller et al., 1992; Miller and Yalamanchili, 1994). The equilibrium and nonequlibrium electrophoretic measurements for borax were found to be consistent and are presented in Table 1. These results provide clear evidence that borax carries a negative surface charge in its saturated brine (pH 9.3), and the sign of the surface charge of borax reverses and becomes positive if the pH is reduced to 8.6. The equilibrium between borax and its saturated brine can be described by the following reaction: [2Na2B407.1OH2O-4Na++B407=+HB4O7 +OH+19H20] It appears that the oxyanions of the borax lattice provide

Jan 1, 1995