Search Documents

By D. E. Healy

In 1989, an autogenous grinding task force was formed to study potential autogenous grinding circuits to replace the existing semi-autogenous grinding (SAG) at the National Steel Pellet Plant in Keewatin, MN. Pilot plant test work, performed by the Hanna Research Center (HRC), indicated that fully autogenous grinding with pebble extraction and crushing could be incorporated into the existing circuits. In 1990, one of the existing 10 grinding mills was converted for an in-plant test incorporating pebble crushing. Full-scale test work were conducted in 1990 and 1991. They verified the pilot plant projections with primary mill throughputs from the fully autogenous circuit using pebble crushing being greater than throughput of the existing SAG mill circuits. The plant produces 5 Mt/a (4.9 million ltpy) of iron ore pellets from 17.6 Mt (17.3 million It) of crude taconite. Crude ore is single-stage crushed by gyratory crushers. Primary grinding is performed by 10 semi-autogenous wet grinding mills. Secondary grinding is performed by five ball mills. Induration is with an Allis Chalmers grate-kiln pelletizing system. The primary grinding circuit consists of 108.2-m-diam x 5.5-m long (27 ft-diam x 18-ft) SAG mills. The mills are charged with 127 mm-diam (5 in.diam) steel grinding balls at from 3% to 5% by volume. Ball consumption rate is about 0.55 kg/t (1.25 lbs/lt) of crude ore feed to the mill. Each SAG mill is powered by two 2.6 MW (3500 hp) motors drawing 5450 kW (7300 hp) at 76% of critical speed. The mills are grate discharge with 16 mm (0.6 in.) slotted grate openings when new. The mill discharge is sized on a trommel screen having 4.5 mm (0.2 in.) openings. The plus 4.5 mm (+0.2 in.) oversize material is returned to the discharge end of the mill by a reverse screw auger located in the center of the trommel. The -4.5 mm (-0.2 in.) trommel screen undersize is pumped to magnetic cobbers.

Jan 1, 1992

By Charles R. Nelson

INTRODUCTION The support of openings in weak sandstones can be achieved by spraying on a liquid grout which will soak in and harden to form a shell. In the St. Peter sandstone of the Minneapolis-St. Paul area of Minnesota compressive strength is increased from about 0.7 MPa (100 psi) to over 7 MPa (1000 psi) to depths of 152 mm (6 in.). A shell of this strength and thickness can pro¬vide support much as shotcrete does but at lower cost. If the spray-grouted shell is inadequate for the support needed, it provides a good base for shotcrete application, which is often difficult in weak or friable sand¬stones because the fresh shotcrete peels off along with a thin layer of sandstone. The spray-grouted technique was initially developed at the Civil and Mineral Engineering Dept. of the University of Minnesota using funds provided by RANN Div. of the National Science Foundation. It was used as temporary support in a 3-m (10-ft) wide, 3700-m (12,000-ft) long storm water drain tunnel by the Minneapolis Sewer Construction Dept. Much of the field development at this construction project was carried out with the cooperation of the personnel of the city of Minneapolis. It also was used for temporary support in a 3.3-m (11-ft) wide, 1800-m (6000-ft) long storm water tunnel built in 1978-1980 for the Minnesota Dept. of Transportation in Minneapolis. In 1975, 2000 m (6500 ft) of 1.5-m (5-ft) wide utility tunnels were spray-grouted for final lining. A 50-m (150-ft) long, 1.2-m (4-ft) wide test tunnel built at the university in 1974 was sprayed for half its length and is being observed for long-term behavior which to date (1981) is good. GROUT MATERIALS The requirements for the liquid grout are more severe than those for injection grouting. Besides being able to penetrate the rock and develop sufficient strength, it must have the following properties: (1) be nontoxic, noncombustible, and have a low odor for spraying under¬ground; (2) have controllable viscosity and setting time for adjusting to the permeability of the sandstone and the required depth of penetration; and (3) must "wet" the sandstone and develop capillary "draw" to penetrate to the required depth. A sodium silicate-based grout with the proper setting agent meets these requirements in the local St. Peter sandstone which has a uniform grain size dis¬tribution curve with D,,, = 0.1 mm, permeability of about 10 to 20 darcys, a porosity of about 25%, and less than 5% of silt size or smaller. It is 97 to 99% pure SiO2. The grout mix consists of Philadelphia Quartz Co. Type N sodium silicate and Celtite 55 Terraset Spray Grade (SG) distributed by Celtite Inc., Cleveland, OH. A volume mix ratio of 100:9:125 for sodium silicate to setting agent to water produces a grout with two to four centipoise viscosity, and initial set at 20°C of 20 min. Setting time is temperature dependent, being longer at lower temperatures, but can be adjusted by varying the setting agent concentration a few percentage points. APPLICATION The application is simple. Spray the grout on the surface until the desired penetration is achieved. Penetration rates of about 5 mm (0.2 in.) per min. are realized in the St. Peter sandstone for up to 152 mm (6 in.) of penetration (30 min. of spraying). The set time must be longer than the spraying time. The spray nozzle should be moved back and forth at a rate that minimizes runoff due to surface buildup. Both a single large hand-propelled nozzle and multiple small machine propelled nozzles have been used successfully. The pumping and mixing equipment used consisted of: (1) a hand-held, hand-pumped, 11-L (3-gal) gar¬den sprayer for test patches (mix by shaking); (2) 76-L (20 gal) batch mixing in barrels with small electric pumps for spraying and mixing; (3) two component pumps consisting of Hypro Model 5300 piston pumps (Hypro Pump Co., St. Paul, MN) on a common shaft driven by an electric motor. The sodium silicate is premixed with a portion of the water and the setting agent is mixed with the rest of the water. The two components are combined and mixed just before the nozzle; and (4) a proportioner metering the setting agent just before the nozzle with the water and sodium silicate premixed and pumped from the surface. The proportioner permits the bulk of the grout to be pumped through a single pipe or hose with only the setting agent stored underground. This is efficient and low in cost. COST The grout mix costs about $0.20/L ($0.75 per gal) in 1978 dollars. Filling all the voids 100 mm (4 in.) deep (25% initial porosity) would have a materials cost of about $5.40/m2 ($0.50 per sq ft). The labor cost of application is usually less than the material cost for normal tunnel jobs. A pump and proportioner cost less than $1000. Sodium silicate storage tanks and pip¬ing costs would depend on the particular site conditions. REFERENCES AND BIBLIOGRAPHY Nelson, C. R., 1977, "Spray Grouting for Tunnel Support and Lining," Underground Space, Vol. 1, No. 3, pp. 241¬ 245. Yardley, D. H., Nelson, C. R., Stocker, T. H., 1974, "So¬dium Silicate Spray Impregnation of Tunnels in the St. Peter Sandstone," Research Report, University of Min¬nesota, 118 pp.

Jan 1, 1982

By Glenn A. Shirey, Robert D. Stoessner, S. Gopalakrishnan

Under a contract awarded by the U.S. DOE-PETC, the Pennsylvania Electric Company and New York. State Electric and Gas Corporation are conducting a test program to evaluate and optimize the performance of a two-inch air-sparged. hydrocyclone (ASH) for cleaning fine coal. The test program is being conducted at the Homer City EPRI Coal Quality Development Center, utilizing Upper Freeport coal classified to minus 0.15 mm as the process feed material. A matrix of experiments has been conducted to establish the influence of significant design and operating variables (including overflow and underflow orifices, inlet slurry pressure, airflow rate, and reagent dosage) on the production of deep cleaned coal. Performance of the ASH is being compared to conventional froth flotation, utilizing advanced techniques and selected flotation reagents. This paper focuses on work completed to-date to optimize the performance of the air-sparged hydrocyclone for cleaning fine coal. INTRODUCTION The Pennsylvania Electric Company (Penelec) and New York State Electric and Gas Corporation (NYSEG) have been evaluating advanced fine coal cleaning processes for the production of compliance grade coal (1.2 pounds S02/million Btu). Laboratory bench-scale froth flotation results were compared with results of advanced flotation processes as performed by various organizations, utilizing the same coal feed. Among the processes evaluated, the air-sparged hydrocyclone (ASH) - developed by Dr. Jan Miller at the University of Utah - was considered to offer the best potential for retrofitting into an existing coal preparation facility because of the increased capacity per unit volume. The ASH has been shown to process over 100 times the throughput of conventional froth flotation cells. In May of 1988, Penelec/NYSEG was awarded a contract from the Department of Energy's Pittsburgh Energy and Technology Center (DOE-PETC) to optimize the performance of a five cm (two inch) ASH for producing compliance grade coal from raw classified minus 0.150 mm (1.00 mesh) Upper Freeport coal, and to compare its performance with conventional froth flotation utilizing advanced techniques and selected flotation reagents. The test work is being conducted at EPRI's Coal Quality Development Center (CQDC) in Homer City, Pennsylvania, utilizing a circuit processing 0.10 to 0.25 tons of coal per hour. Other members of the Penelec/NYSEG contract team include the University of Utah, Advanced Processing Technologies, Inc., Davy Dravo, Kaiser Engineers, Ebasco Services, and Management and Technical Systems. ASH DESCRIPTION AND OPERATION The ASH-2 System is a five cm ID (actually 4.76 cm) air-sparged hydrocyclone manufactured by Advanced Processing Technologies, Inc., under license from the University of Utah, for the purpose of research, evaluation and development. A schematic drawing of the air-sparged hydrocyclone is presented in Figure 1. Basically the air-sparged hydrocyclone consists of two concentric right-vertical tubes, a conventional cyclone header at the top, and a froth pedestal at the bottom. The inner tube is a porous tube through which air is sparged radially. The outer nonporous tube simply serves as an air jacket to provide for even distribution of air through the porous inner tube. The slurry is fed tangentially through the conventional cyclone header to develop a swirl flow in the radial direction. Air is sparged through the jacketed, inner porous tube wall and is sheared into small bubbles by the swirl flow. Hydrophobic particles in the slurry collide with these bubbles and, after attachment, are transported radially into a froth phase which forms on the cyclone axis. The froth phase is stabilized and constrained by the froth pedestal, thus moving towards the vortex finder of the cyclone header and discharged as an overflow product. Hydrophilic particles generally remain centrifuged in the slurry phase and are discharged as an underflow product through the annulus created by the froth pedestal.

Jan 1, 1989

By Joseph L. Walker

INTRODUCTION In looking over the diversity of the papers that are being presented during this meeting on The Industrial Practice of Fine Coal Processing one cannot help but be impressed with the progress the industry has made in beneficiating fine raw coals, and in the number of innovations that are under study to increase recovery and to improve the quality of the finest size particles produced at a preparation plant. Those who are not in our industry, who look at the technical program, might be confused, when they see that the papers include such topics as cleaning to zero and heavy media cyclone cleaning 28M x 100M; application and performance of fine coal jigs, tables, hydrocyclones, spirals, and flotation; filtration, dewatering and thermal drying; and ancilliary aspects such as fines handling, instrumentation and process control plus screening and liberation. This is an observation and not a criticism of the diversity of papers being presented. The program fits within the parameters of fine coal preparation as the term "coal preparation" is defined, but perhaps not as the term "fine coal" is defined. Admittedly these definitions are sometimes loose. However, each should be made more definitive if a meaningful fine coal preparation overview is to be possible. COAL PREPARATION DEFINED In the opinion of the writer, the term "coal preparation" should be defined as including all work activities that are performed on all components of a run-of-mine coal from the time it leaves a surface reserve or an underground portal, until all components are either ready for shipment or are stowed on, or in the earth in an environmentally acceptable manner. This definition excludes the work that is performed to improve coal quality or handleability inside the mine or pit. This work is commonly called "face" or "pit" preparation. FINE COAL DEFINED The term "fine coal" means different things to different people. In the past, anthracite finer than No. 3 Buckwheat (3/16 in.) was considered fines which often were discarded from the breaker as an unsalable by product; and bituminous coals finer that about 1/4 in. (slack) were either discarded or sold at cost to prevent an accumulation of "no bills" which referred to a product having no immediate sale. Although a standard definition of "fine coal" is yet to be accepted by the industry, a designation of coals by particle-size range has evolved through the wide use of three-circuit flowsheets, wherein the optimum size range by washer type determines the size fractions: a coarse-coal circuit for the plus 3/8 in., a -11- coal circuit for the 3/8 in. x 28 Mesh, and a fine-coal circuit for the remaining 28 Mesh x 0. Logically, the term "fine coal" should refer to some resultant size fraction or some top size by zero. This definition should include a 28 Mesh x 0 feed to a flotation cell but should exclude a 28 x 100 Mesh size fraction fed to a heavy media cyclone. Failure to include the resultant sizes in the definition precludes accounting of the full range of particle sizes present in the feed to the plant. Since the 28 Mesh x 0 fraction is treated in the final circuit of the three discreet washing circuits, now almost always used in U.S. commercial applications, and since it contains all of its resultant sizes, in the writer's opinion, it should be designated "fine coal". FINE COAL CLEANING Cleaning coal fines, which in the past was at best marginally profitable, is now an essential part of coal preparation strategy for a number of reasons : 1. Raw coals now cost more to produce and

Jan 1, 1988

By David M. Stearns

Introduction Millions of tons of coal are shipped by rail each year in the US. Loading those trains efficiently is a topic being studied by coal producers and railroads. Alternatives range from volumetric loadouts to systems that weigh the product, using either belt conveyor scales, track scales, or weighbins. Each system has its proper application and each has pros and cons, depending on the volume, nature, and layout of the specific loading facility. In recent years, a batch weighing system has been developed that offers significant benefits to higher volume operations shipping by a unit train. The Unitrain Loadout System (ULS) features a precision electronic batch weighing system married to a mechanical loadout system by a microprocessor-based digital control system. A batch weighing loadout could be designed using analog electronics or even a mechanical weighbin. But it is the adaption and use of a microprocessor that provides the primary benefits of a unit train loading facility. Those benefits are accurate weight determination, optimum car use, overload prevention, speed in loading, optimum labor use, and documentation of the loading process. Unitrain History The first ULS was erected for iron ore applications in Canada during the 1960s. The first domestic system on coal was installed at Pittsburg and Midway Coal Co.'s McKinley mine at Gallup, NM, in 1978. At present, there are more than 20 Unitrain Loadout Systems in operation worldwide. They are loading minerals, concentrates, and coal. Eleven systems are presently in service or under construction in the US. In the last several years, a number of comparable systems have also been installed. As a result, the concept has been well proven and is now commercially available from several sources. Unitrain Operation The basic operation of a ULS involves loading a unit train with individually weighed batches of coal while it moves slowly under a loadout tower. Typical coal systems are designed to load at rates of 2.7 to 6.3 kt/h (3000 to 7000 stph). The contents of each car can be accurately weighed. This meets the requirements of the National Bureau of Standards Handbook No. 44 for static weighing. Also, cars can be loaded within very close tolerances. This optimizes railcar usage and prevents overloads. Principal components of the loadout system are the main feed conveyor, the surge bin, surge bin gates, weighbin, weighbin gate, flood loading chute, control room, calibration weights, and hydraulic and electrical systems (Fig. 1). Main feed conveyor: Coal is delivered from a storage area by the main feed conveyor. The conveyor fills the surge bin. It should be capable of conveying coal at the same rate at which it is being loaded. Usually the feed conveyor is equipped with a conveyor scale that can display rate in the control room and control feed to the belt. The surge bin will typically store 227 to 272 t (250 to 300 st) of coal. If the loadout is installed in an over-the-track silo, there is no need for a surge bin. Four double-bladed, hydraulically actuated gates are incorporated into the bottom of the surge bin. These gates open to fill the weighbin. Automatic pre-act points are selected by the control system. They shut these gates as the desired weight is approached. The final amount of coal is added by a single gate that closes. This results in a weighed batch of coal within ± 0.5% of the selected optimum load per railcar. Thus, if loading 91-t (100-st) capacity cars, the system is set to batch up 90.2 t (99.5 st) batches. Each car is expected to be loaded between 90 and 91 t (99 and 100 st). Weighbin: The weighbin is next in the flow scheme. Typically, it is

Jan 3, 1985

By P. D. Chamberlain

Agglomeration for the heap leaching of gold and silver ores is the process of attaching fine ore particles-less than 150 µm (100 mesh)-to coarser particles. The fines are thus immobilized and uniformly distributed throughout the heap. This is needed to make heaps more porous and uniformly permeable to the flow of a leaching solution. This article describes liquids and binders used to form agglomerates. Agglomeration techniques and equipment are also described. These include drum and disk agglomerators, stockpile and belt agglomeration, and the use of vibrating deck and reverse-running conveyors. The need for agglomeration The main reason for agglomerating heap leachable ores is to prevent percolation problems in a heap. This is different than pelletizing iron ore. In the latter case, the purpose is to bond very finely ground material into coarser, strong particles. These particles can then withstand the rigors of transportation and the loading forces in a blast furnace. Most percolation problems in heaps are caused by the segregation of coarse and fine particles during heap construction. This segregation creates areas with significantly lower permeability because fines plug the channels between coarse particles. Consequently, the leach solutions are subject to channeling, following the path of least resistance. They percolate downward through the coarse ore areas and bypass or barely wet areas that contain large amounts of fines. This results in lower extraction, longer leach time, and higher reagent consumption. Permeability problems in heap leaching are compounded because of the vertical migration of fine particles after the heap has been built. These fines eventually accumulate in pockets or layers and impede uniform solution flow. The fines move because they are not attached to a coarser particle. Moreover, the sluicing action of leach solution flowing over the surface of particles may break the bond between fine and coarse particles. Some fines are also created by the decomposition of ores after being wetted. Percolation problems are minimized if the fines can be attached to the coarser particles. Fines are thus uniformly distributed and rendered immobile within the heap. Heinen (1979) conducted comparative tests of nonagglomerated and agglomerated gold and silver ores. He showed that percolation rates through agglomerated material usually are improved by 10 to 100 times. This reduces leach time for economic recovery to less than one-third of that for nonagglomerated material, and total recovery may be increased. Cyanide consumption is also reduced in most cases because of the shortened leach time. Even run of mine (ROM) ore should be agglomerated. This can be achieved with moderate success by passing each truckload of ore under a water spray to bring total moisture of the ore up to about 8% to 10%. The ROM ore will mix when the ore is dumped, when it cascades down the slope of a heap, or when it is pushed up with a dozer. Overall, agglomeration allows low-cost heap leaching technology be applied to gold and silver ores and tailings that could not be economically processed by conventional methods, according to Phariss (1982). How agglomeration works Agglomeration means that the clays and fine particles contained in the ore adhere to the coarser particles. This creates a coating of fines around the coarse particles. When the ore has few fines and the percolation problems are not severe, the addition of a liquid may be all that is necessary to make the few fines adhere to coarser particles. However, when the quantity of fines is substantial, e.g., +10% -75 µm (-200 mesh), a binding agent may be needed. Use of solution only for agglomeration There are many forces that cause agglomeration and one of them is the surface tension of water. Water-saturated particles collide with each other or with

Jan 12, 1986

By Floyd C. Bossard

A method of hand-calculating the air-cooling requirements of a conceptualized underground mining operation is presented for the reader's orientation. Separate air heat load calculations were conducted for adiabatic compression, electromechanical equipment, wallrock, broken rock, groundwater, and blasting operations. The air heat sources were calculated for four mining levels under conditions representative of a typical planned mining operation at depth. The total heat gain on each level will approximate the level air-cooling requirements. The results of these hand calculations can be further modified by the use of mine ventilation computer software that refine the heat-source calculations, predict underground ambient air temperatures, and establish the air-cooling requirements of a mine. INTRODUCTION The principal mining method employed is a modified vertical crater retreat (VCR) blasthole operation. Typical scopes range from 20 to 40 feet wide, 75 to 100 feet long, and 100 feet in height. Six and one-half inch blastholes are drilled with an "in the hole" hammer drill. Eight-foot high horizontal rounds are blasted down. Mucking of ore from the undercut to orepass is done with LHD equipment operated from a remote control station. Backfill includes hydraulically placed tailings with cement, and waste rock when available. Stope access is from ramp sub-levels on 50-foot vertical intervals. Crosscuts are ramped down to the first stope cut 25 feet below the sub-level elevation. Then the crosscuts are raised by taking down the back when each stope cut is completed, until an elevation 25 feet above the sub-level is reached. The crosscuts are filled with tailings, and/or waste rock. See Figure 1. Typically, a two-pronged approach to defining the air-cooling requirements is conducted. First, the principal sources that make up the air heat load are individually hand-calculated. Second, projections of mine heat load are calculated by utilizing computer modeling techniques. This paper discusses the first method (hand calculations) of determining the individual components of heat flow into the mine. CALCULATED AIR HEAT LOADS Adiabatic Compression The plans for the mine include delivery of 300,000 cfm of air to ventilate the lower level operation. This is equivalent to approximately 200 cfm/ton of rock produced (300,000 cfm/1500 tpd of ore and waste). As air flows down a shaft, with no heat interchange between the shaft and air and no evaporation of moisture taking place, the air is heated in the same way as if it were compressed in a compressor. Dry air increases in temperature about 5.4°F per 1000 feet. One BTU is added to each pound of air for every 778 feet of decrease in elevation, or is subtracted for the same elevation increase. For dry air, the dry-bulb temperature change is 1/(0.24 x 778) = 0.00535°F/ft., or 1°F/187 ft. elevation. Auto-compression may be masked by the presence of other heating or cooling sources, such as shaft wallrock, groundwater, air and water lines, electrical facilities, etc. The major factors influencing the temperature of the air delivered underground by a shaft are (1) the night time cool air temperature's effect on the rock or lining of the shaft, (2) temperature gradient of ground rock related to depth, and (3) evaporation of moisture within the shaft which increases the latent temperature and decreases the sensible temperature. [Calculation For Adiabatic Heat of Compression] a. Assumptions: 1. Three hundred thousand cfm of fresh air at 3000 level has increased in temperature during the summertime to the point where it has little available cooling power. Air-cooling will be required on 3000 Level and below. 2. Elevation of 3000 Level is approximately +1000 ft. above sea level.

Jan 1, 1993

By John E. Dreier

The Pachuca-Real Del Monte district, Hidalgo, Mexico (Figure 1) has produced roughly 1,500,000,000 oz Ag., 6,000,000 oz Au. and minor Pb., Zn, and Cu from about 100,000,000 tons of ore since discovery early in the 16th century. Production has come from epithermal vein deposits that have been up to 8 km in length, 650 m in vertical extent and 45 m in width but which averaged 200-400 m in length and vertical extent and 2-4 m in width. Ore deposits occur through-out an area about 12 km E-W by 10 km N-S (Figure 2) and between the elevations of 2000 and 2800 m above sea level. The deposits occur as shoots in the veins and terminate laterally where the veins narrow, upward where the veins pinch out and downward where the veins decrease in silver con-tent to below the cut-off grade; the veins are known to extend 500 m below the bottoms of the ore bodies and probably extend to much greater depths. The district lies wholly within a mid-to-late Tertiary volcanic rock sequence that is made up principally of andesite and dacite lava flows and flow-breccias. Although the bottom of the sequence is not exposed in the district, stratigraphic constructions indicate the volcanic rocks to have an aggregate thickness of 2000 to 2500 m. Structurally, the area is characterized by gentil warping and normal faulting of the volcanic rocks. Pre-ore modifications to the volcanic rock sequence consisted mainly of faulting along west-trending, south-dipping normal faults and north-trending high-angle faults.

Jan 1, 1978

By Aditya Rana, A. K. Mishra, M. M. Singh, N. K. Bhagat

Many cases of derailments, train accidents and traffic interruptions due to the slope failure and rockfall have been reported in the past on Konkan Railway trackline because of unstable rock slopes. The authorities installed stability enhancement measures such as wire netting, rock bolting and shotcreting which were found insufficient. The failed geotechnical measures prompted authorities to excavate the unstable slopes. Rock blasting is generally not preferred for the excavation of slopes along the transportation route. Stabilisation and widening of slopes by blasting need closure of the route for a considerable time. Discontinuing traffic on Konkan Railway route was undesirable as it connects two important port cities Mumbai and Mangalore. The paper presents an innovative directional controlled blasting (DCB) technique for excavating three unstable slopes (i.e. cuttings) in hilly terrain without disturbing the traffic. New empirical formulae for small geometry blasting to estimate burden and throw have been developed. A unique sequence of firing and excavation with powder factor varying between 0.04 and 0.39 kg/m3 irrespective of the actual free face was used for rock blasting. The process controlled the throw of blasted material over the track and resulted in a smooth and stable wall. The slope angle has been changed from 80 to 82° to 45–47° and probability of wedge, planar and toppling failures was eliminated. About 100 such accident-prone slopes (52 km cumulative length, 7% of total route) situated at discrete locations along Konkan Railway route were stabilised using the technique. No incident of rockfall and slope failures along these slopes has been observed after the stabilisation.

By Robert E. Kirby

The potash mine and refinery of AMAX Chemical Corporation is located 26 miles east of Carlsbad on the north end of the potash producing areas. This property was developed between 1949-1952 and production of muriate began in August 1952. Two shafts, located 400 feet apart, were sunk to the ore zone 900 feet below the surface. A 15-foot diameter ore hoisting shaft equipped with a 1000-ton ore pocket, crusher and automatic skip loading equipment plus a 20-foot diameter service shaft with a curtain wall for exhausting air were provided. Sylvinite production to date has been from the first ore zone, stratigraphically, the lowest of ten identified mineralized zones in the salt section. This first ore zone has been the top producer of Muriate of Potash over the years. In our area we do have some very marginal mineralization in the third ore zone some 35 feet above the first. We have not mined from this seam as yet but we are investigating the economics very closely with access slopes and an experimental mining panel. Our ore seam is generally flat lying but does have local gentle rolling with grades from 2% to 5%. The mine was designed and developed from the beginning with conventional mining practices and rail haulage from the mining section to the ore pocket. As mining progressed in the early years, production increased from less than one million tons annually to a maximum of 3,935,000 tons in 1964. Efficiency also kept pace with the increasing production and continually improved until the efficiency of the production crews reached an average of 100.3 tons per manshift in 1964.

Jan 1, 1974

By J. R. Patts, A. B. Cecala, E. J. Haas

Exposure to respirable crystalline silica (RCS) remains a serious health hazard to the US mining workforce who are potentially exposed as various ore bodies are drilled, blasted, hauled by truck, crushed, screened, and transported to their destinations. The current Mine Safety and Health Administration (MSHA) permissible exposure limit (PEL) for RCS remains at approximately 100 μg/m3 , but it is noteworthy that the Occupational Safety and Health Administration (OSHA) has lowered its PEL to 50 μg/ m3 (with enforcement dates staggered through 2022 for various sectors), and the National Institute for Occupational Safety and Health (NIOSH) has held a 50 μg/m3 recommended standard since 1976. To examine a method for reducing RCS exposure using a NIOSH-developed video exposure monitoring (VEM) technology (referred to as Helmet-CAM), video and respirable dust concentration data were collected on eighty miners across seven unique mining sites. The data was then collated and partitioned using a thresholding scheme to determine exposures that were in excess of ten times the mean exposure for that worker. Focusing on these short duration, high magnitude exposures can provide insight to implement controls and interventions that can dramatically lower the employee’s overall average exposure. In 19 of the 80 cases analyzed, it was found that exposure could be significantly lowered by 20% or more by reducing exposures that occur during just 10 min of work per 8-hour shift. This approach provides a method to quickly analyze and determine which activities are creating the greatest health concerns. In most cases, once identified, focused control technologies or behavioral modifications can be applied to those tasks.

By J. R. Patts, A. B. Cecala, E. J. Haas

Exposure to respirable crystalline silica (RCS) remains a serious health hazard to the US mining workforce who are potentially exposed as various ore bodies are drilled, blasted, hauled by truck, crushed, screened, and transported to their destinations. The current Mine Safety and Health Administration (MSHA) permissible exposure limit (PEL) for RCS remains at approximately 100 μg/m3 , but it is noteworthy that the Occupational Safety and Health Administration (OSHA) has lowered its PEL to 50 μg/ m3 (with enforcement dates staggered through 2022 for various sectors), and the National Institute for Occupational Safety and Health (NIOSH) has held a 50 μg/m3 recommended standard since 1976. To examine a method for reducing RCS exposure using a NIOSH-developed video exposure monitoring (VEM) technology (referred to as Helmet-CAM), video and respirable dust concentration data were collected on eighty miners across seven unique mining sites. The data was then collated and partitioned using a thresholding scheme to determine exposures that were in excess of ten times the mean exposure for that worker. Focusing on these short duration, high magnitude exposures can provide insight to implement controls and interventions that can dramatically lower the employee’s overall average exposure. In 19 of the 80 cases analyzed, it was found that exposure could be significantly lowered by 20% or more by reducing exposures that occur during just 10 min of work per 8-hour shift. This approach provides a method to quickly analyze and determine which activities are creating the greatest health concerns. In most cases, once identified, focused control technologies or behavioral modifications can be applied to those tasks.

By Joseph E. Schiller

As part of the Bureau of Mines research on identification of asbestos fibers, a method was discovered for removing asbestos fibers from water. This report describes this process and its theory of operation. The surface charge of amphibole minerals in potable water was found to be negative, and in common with most other silicate minerals, the isoelectric point of amphibole asbestos fibers lies in the pH range 4 to 6. Magnesia is one of the rare oxides with a high isoelectric point of about 11. The surface of magnesium oxide will, therefore, be positively charged in water at pH 6 to 6, while the electric charge an. amphibole asbestos fibers is negative. Because of the electrical attraction of these oppositely charged solids, it was possible to remove negatively charged asbestos fibers from water by percolating water that contained suspended-asbestos through a porous bed of magnesium oxide particles. A bed of 100/200-mesh MgO was added to a glass tube with a cotton plug. Aqusous suspensions of 1 to 10 ppm cummingtonite asbestos (amosite), crocidolite, and chrysotile were percolated through this porous plug of MgO. Even though chrysotile is positively charged below pH 11, MgO is an effective filter medium for removing it. The highly basic MgO appears to cause the chrysotile to flocculate and attach to particles within the bed. Scanning electron photomicrographs indicated that essentially complete removal of the fibers from water was accomplished. Other materials tested as contact filter media and found to be ineffective or only partially effective were sand, calcite, carbon, diatomaceous earth, and cellulose. Acidic alumina collected amphibole fibers but not chrysotile fibers.

Jan 1, 1980

By James A. Clum

The shift to lower grade iron ores, e.g., magnetic taconites, resulted in a shift towards agglomeration practices which could effectively deal with the fine particle sizes generated in the liberation and concentration processes. If the iron bearing particles are not extremely fine, e.g., -1/2 in + 100 mesh, sintering is an adequate means of agglomeration.(1) For the more dispersed ores which result in fine concentrates (more than 80%,-325 mesh), the fine particles must be agglomerated by mechanical means before final hardening, e.g., pelletizing.(1) In the pelletizing process, the essentially mechanical step of balling is extremely critical since the strength of the pre-fired pellet determines the firing conditions. The ability of the proper amount of moisture to hold together a ball of the very fine ore concentrate particles has long been recognized. It is a charac¬teristic of the reaction at the interface between the ionic solid of the ore concentrate and the polar water phase which overlays the solids. The strength in such moisture laden balls, the green strength, is a strong function of the water content (Figure 1) and thus deteriorates rapidly for either a small excess or a small deficiency of moisture. For purposes of commercial practice, the control of green ball handling properties by accurate control of ball moisture is too demanding. Moreover, as the ball is heated towards its hardening tempera¬ture, e.g., 1300'C, there is a substantial drying reaction which would lead to such a low dry strength of the ball that it would not support its own weight much less the weight of other balls in a firing furnace.

Jan 1, 1976

By Neal P. Cochran

My talk today is an effort to project our thinking forward a few years to that point in the future when we will be producing gas and liquid fuels from coal. To those of you that are concerned with meeting today's production quota--my remarks may sound like a Buck Rogers projection. Not so. Although I'm talking about the future, please understand that my thoughts are focused on the near term future. The coal plant of the future will be an energy plant producing energy products to meet the needs of our modern society. This plant is an outgrowth of Government- industry cooperation. Much of the research work that makes such a plant possible was jointly funded, other work was a continuation into development of private research expenditures, other work was performed by private research organizations with Government funding, other work was performed in Government laboratories, both State and Federal. As indicated above, and in the title of my talk, I expect to describe the coal plant of the future. I will describe a plant producing Carbon, Oil, Pipeline gas, and Electricity--Project COPE. This plant copes with the Atomic-Nuclear Group. As stated above, this plant is an energy plant. Energy plants today are large-scale operations. Power plants of 5,000 megawatts are projected, refineries of 500,000 daily barrels are near term probabilities, 48" transcontinental gas pipelines are with us today and 25,000 daily ton mines are here as well. The COPE plant will be of the same general size and require 50,000 - 100,000 tons of coal per day.

Jan 1, 1969

By Ted C. Mathews

Pre-concentration by sorting is as old as the mining industry. Inevitably waste rock is broken with ore. Separation may take place in the stopes by selective mucking and transport, or selective loading and transport of ore and waste by 12-cu, yd. shovels and 100-ton trucks in large open pit operations. Hand sorting of 1" to 16" pieces, formerly widely practiced at underground chutes and prior to or following first stage crushing, has greatly diminished in the United States with its high per hour labor cost, although still widely practiced in other countries and, surprisingly, in some of the most modern mills where inexpensive labor is available. The economics of pre-concentration by sorting is attractive since elimination of waste at any cost less than the total milling cost will show a profit. The size at which sorting should be under- taken is determined by the size at which liberation is secured to yield a product of ore and waste, and greatly influences cost and the point in the flowsheet where sorting can be undertaken. In some instances mining costs can be reduced where an inexpensive sorting system to eliminate waste is combined with a less selective and less costly mining method. This is particularly true in deep mines where hoisting capacity determines the rate of extraction and sorting is performed underground, with waste utilized for fill. Hand sorting, while efficient, is expensive because of low sorting rates. Peale (1) quotes man hour capacity at 1/3 ton per hour on sizes 3/4" to 3" copper ore, and up to 5 tons per hour on rich iron ore sized from 4" to 16".

Jan 1, 1974

By A. D. Davis

Naturally occurring arsenic has been detected in water from inactive mines and several wells in the Keystone area of the Black Hills. Keystone City Well #4 showed a total arsenic concentration of 69 parts per billion (ppb) during pumping in 2004. The well is not used currently except for arsenic-removal research. Water also can be pumped from a mine shaft in the abandoned Keystone ? Columbia ? Bullion gold mines, whose interconnected workings flooded after mining ceased in the 1930s. This water has been considered for local supplies, although it would require treatment or blending because of arsenic concentrations. According to historical records, the Bullion Mine produced 100,000 pounds of white arsenic. An aquifer test on the mine workings was performed at a pumping rate of about 550 gallons per minute. The arsenic concentration in pumped water was 33 ppb. In addition to wells within Keystone, samples from the water supply at nearby Mount Rushmore National Memorial have shown arsenic concentrations greater than 10 ppb, the U.S. Environmental Protection Agency?s maximum contaminant level for arsenic in drinking water. Limestone-based technology for removal of arsenic from drinking water is an innovative and promising method that has been considered for treatment of water at Keystone. In batch testing with water from Keystone Well #4, limestone removed about 65% of arsenic from solution. A related problem of arsenic removal from drinking water involves disposal of the waste product, which is critical for commercial viability of removal technologies. Current research is investigating thermal stability of the waste in order to determine the potential for recycling of the material during manufacture of cement.

Jan 1, 2010

By William E. Horst

Ore grindability data based on various techniques to measure comminution characteristics are used to evaluate the crushing and grinding efficiency in mineral processing operations. The importance of achieving improved comminution efficiency, in terms of energy consumption, has been emphasized by recent increases in the cost of electricity. From a theoretical standpoint, the objective of a grinding operation might be expressed as maximizing the production rate at a given grind size while minimizing the energy consumption per ton of material ground. To achieve this objective in an industrial operation presents a challenge especially in view of changes in the raw material which are normally encountered. Various procedures1-5 used to determine energy-size parameters and grindability measurements have been reported. The proposed relationships between particle size and energy consumed have much in common between many of the techniques for measuring grindability. Probably the most widely used parameter to measure ore grindability in recent years is the Bond work index6, Wi, which is defined as the kw-hr per ton required to break from infinite size to a product size of 80 percent passing 100 microns. If the breakage characteristics of a material remain constant over all size ranges, then the calculated work index would be expected to remain-constant since it expresses the resistance of material to breakage. However for most naturally occurring raw materials differences exist in the breakage characteristics depending on particle size, which can result in variations in the work index depending on grind size. Consequently work index values are generally obtained for some specified grind size which typifies the comminution operation being designed or evaluated.

Jan 1, 1975

By Arthur B. Cummins

It is appropriate to discuss diatomite at an AIME meeting in San Francisco. California has led the nation, and the world, in the production of diatomite products for more than 40 years. It appears that our State will maintain this position for a long time to come. Diatomite today ranks high in the list of California mineral commodities, with a total value since 1889 well in excess of $100 million. California's production has more than doubled since 1942. In 1959, the value of diatomite production in California is exceeded only by the following industrial mineral materials-borates, stone, sand and gravel and cement20,28. Also, not unimportant is the production of diatomite in each of our other Pacific Coast States, including Nevada. First, let us define diatomite. It is a sedimentary siliceous rock of organic origin, which consists mainly of the fossilized remains of diatoms. Diatoms are microscopic aquatic plants which have the capacity of precipitating silica bio-chemically so as to form a part of their organized structure. In the economic sense, we are not concerned with the numerous types of sedimentary materials which contain minor amounts of diatom residues, or with impure or highly indurated siliceous materials, which may have been formed from diatomite2. Diatomite (synonyms--diatomaceous earth, kieselgurO is a term generally restricted to material of such qualities as to be suitable for commercial uses. Diatomaceous silica is a designation which is best used to distinguish the basic component of diatomite, that is, the substance of the silica shell itself, and which is the major' constituent of beneficiated diatomite or processed diatomaceous products.

Jan 1, 1959

Ground falls (roof and rib) have historically been responsible for about 50% of all fatalities in bituminous underground coal mines. In some recent years, the number of ground fall fatalities has approached zero, indicating that significant progress has been made. On the other hand, last year?s Crandall Canyon twin mine disasters, in which 9 miners perished in violent coal bumps, provided a stark reminder that complacency is premature. One important success has been a great reduction in the number of miners killed inby roof supports. Throughout the 1990?s these accounted for nearly half of all roof fall fatalities, but there have just been two inby incidents in the past 5 years. Progress has also been made in pillar recovery. Safer mining technologies have reduced the number of fatal incidents, but some recent failures have underlined the need for careful management of the retreat mining process. On the other hand, more than 400 miners continue to be injured each year by rock falling from between supports, and 100 more are injured by rib falls. Together, these two categories also account for a large percentage of recent ground fall fatalities. Available technologies such as roof screen, rib bolting, and inside control roof bolters could dramatically reduce injury and fatality rates if they were used more widely. Further advances in these areas will likely be the next big advance in ground control safety. Disclaimer: The findings and conclusions in this report have not been formally disseminated by the National Institute for Occupational Safety and Health and should not be construed to represent any agency determination or policy.

Jan 1, 2009