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By R. A. Matuszak

Historically most mine management teams have been forced to operate in a more or less subjective manner. The objectives of mine management are simple and clear. Like other operational managements, mine management teams have the traditional well defined goals of improving efficiency and production while decreasing costs. Unlike other operational managements however, mining teams deal regularly with a vast assortment of uncertainties caused by: - Key production tools being mobile and widely dispersed, and - Non-uniform mining conditions At best in most operations these conditions cause obstacles to productivity and at the same time cause long delay in managment awareness of the obstacles. Planning with delayed feedback of information about necessary deviations is incomplete planning. For these reasons increased operational efficiency remains a constant though somewhat elusive goal.

Jan 1, 1985

By Ross W. Smith

Many studies have been made of the microflotation of quartz (1,2,3,4,5,6) and alumina (3,5,6, 7,8) using various anionic and cationic hetero- polar hydrocarbon surfactants as collectors. A few studies have been made on these minerals using amphoteric hydrocarbon collectors ( 9,10, 11,12) and only a few studies have been made on the use of fluorocarbon surfactants as flotation collectors (13,14). The present investigation compares microflotation response of the two minerals in hydocarbon and fluorocarbon collector systems of the anionic, cationic and amphoteric types. An attempt is made to relate flotation response of the minerals to their electrokinetic behavior and the pKa values of the collectors.

Jan 1, 1979

By Henry O. Ash

On June 16, 1982, Rio Blanco submitted a request to the US Minerals Management Service for a suspension of operations under its federal oil shale lease. The company also wanted a five-year delay in the requirement for further minimum royalty payments and a five-year extension of the primary lease terms of 20 years. The request was granted in late July in the midst of yet another slump in oil shale activity. Rio Blanco holds one of four Federal oil shale leases that were granted under the Interior Department's Prototype Oil Shale Leasing Program in 1974. That year the DOI offered for lease six federal oil shale tracts, two each in Colorado, Utah, and Wyoming. The leases in Colorado and Utah were granted for substantial bo¬nus bids as shown in Table 1, but the Wyoming tracts had no takers. In the eight years hence, no oil shale has been produced commercially. Indeed, the prospects for successful oil shale development have fluctuated widely, continuing a 60-year- old cyclical pattern in the fledgling industry. The prototype program was developed and initiated while this infant industry struggled to begin anew, without the benefit of any federal oil shale leasing regulations, or any history of federal oil shale leasing at all. Permanent regulations were not required because the effort was limited and designed to provide the foundation for any future leasing program. Section 21 of the Mineral Leasing Act of 1920, a Nov. 30,1973 Federal Register notice, and the leases themselves constitute the legal and regulatory framework for leasing and developing prototype oil shale tracts. The prototype leases contain some unusual provisions including installment payments of the bonus bids, credit of early development costs against payments due the government, adjustment of royalty rates according to regional crude oil price fluctuations, and unusually extensive and detailed environmental stipulations. Several of the provisions were intended to provide incentives for early development and financial penalties for failure to develop. The requirement for minimum royalty payments in lieu of actual production royalties beginning in the sixth year of the lease falls in the latter category. After the sixth year, the required minimum payment is increased east year through the 15th year by an increment equal to that due in the sixth year. The minimum royalty then remains the same through the final five years of the primary lease term. Minimum royalty is calculates on a different production level for each prototype lease, based on the quantity of 125 L/t (30 gal/st) shale estimated to be recoverable from. each tract by room-and-pillar mining. That mining system was selected for reference as the only proven extraction method for oil shale at that time (early 1970s) The minimum royalty (production) rate was calculated for each tract by direct comparison of its estimated recoverable 125-L (30-gal) resource to a hypothetical tract containing 1.9 Gt (2.1 billion st) of recoverable 125-L (30-gal) shale. For the hypothetical lease tract, the minimum production for the sixth year was calculated to be 3.2 kt/d (3,500 stpd), increasing by a like amount each lease year through the 15th to a daily production rate of 31.7 kt (35,000 st). Under this formula, the hypothetical lease tract would yield $16 million in minimum royalty payments over the 20-year primary lease term if there were no production, no development credits taken, and no adjustment of the royalty rate due to fluctuation of regional crude oil prices. The production rate of 31.7 kt/d (35,000 stpd) was a somewhat arbitrary figure that was considered, in the early 1970s, to represent a minimum size for a commercial oil shale operation. A production rate of 7.9 dam3/d (50,000 bbls/day) soon supplanted that smaller mine production figure as a rule of thumb for the minimum size for a commercial facility. The proportional relation-

Jan 11, 1982

By Ronald B. Gabitzsch, Norbert P. Hanke

During 1982, record advances have been achieved in the excavation of vertical shafts with a V-Mole (a vertical shaft boring machine). The V-Mole, supplied by Wirth GmbH in Erkelenz, West Germany, has been used in West German coal mines since the early seventies, mainly for the construction of internal shafts. In 1981, it was brought into the United States for the construction of a series of ventilation shafts with depths varying between 512 m (1680 ft.) and 622 m (2040 ft.) excavated to 7.0 m (23 ft.) diameter. These shafts were needed for the development of coal mines for Jim Walter Resources, Inc. in Brookwood, Alabama. After a description of the mechanized shaft excavation system with the V-Mole, the conditions and parameters, such as ground, shaft dimensions, and linings are discussed. These factors are important considerations for a technically and economically feasible utilization of the V-Mole. In addition, the advantages and disadvantages of using the V-Mole technology in comparison to other shaft construction methods are evaluated.

Jan 1, 1983

By Fred N. Kissell, George H. Jr. Schnakenberg

INTRODUCTION A variety of instruments are available for measuring or monitoring the performance of underground mine-ventilation systems. In general terms, the instruments may be classified as those that measure air velocity and those that measure gaseous concentrations. All costs herein are in terms of 1978 US dollars. The mention of a specific manufacturer or device is not intended to be an endorsement by the US Bureau of Mines. AIR-VELOCITY INSTRUMENTS The basic instruments used for measuring the air velocity in mines are the vane anemometer and the smoke tube. Vane Anemometer Of the air-velocity instruments, the 102-mm (4.0¬in.) vane anemometer is the most common and is available as either a low- or high-speed type. The low-speed anemometer generally is the most suitable for measuring the velocities in ordinary airways. For a rough check of the velocity in an airway, it usually is satisfactory to hold the anemometer by hand, positioning it in the center of the airway for 30 sec. However, the resultant error may be as high as 25% , and such a hand-held approach is unsuitable when accurate or reliable measurements are required. To obtain more accurate measurements, the proper procedure is as follows: 1) Since holding the anemometer by hand generally causes the instrument to read about 15% high, it is mounted on a 0.6-m (2-ft) extension rod. 2) The airway is divided into equal right and left halves. A 1-min traverse is used in each half, moving the anemometer smoothly up and down in a zigzag pattern so that the entire half is covered within the allotted minute. 3) The manufacturer's correction table is applied to the readings to adjust the velocity calculation as necessary. Whenever possible, anemometer readings should be obtained in a long straight section of airway that has a constant cross-sectional area. Bends and obstructions should be avoided, since they cause turbulence and other discontinuities in the airflow and can degrade the accuracy of the velocity measurements. Although a series of velocity measurements at one location usually corresponds to within a few percent, this is not an indication that the airflows calculated from those readings are completely accurate. One reason is that the correction table provided with the instrument generally is not for that specific instrument; instead, it represents the average correction for all such instru¬ments made by the particular manufacturer. Most cor¬rection tables specify a correction factor of from 0 to 15%, depending upon the velocity. However, even after correction, the instrument error still may range from 3 to 5%. At low velocities such as those below 0.76 m/s (150 fpm), the instrument error can be two or three times greater than this, ranging from 6 to 15%. The new ball-bearing anemometers generally perform somewhat better at low velocities than did the older conventional anemometers. Another source of error is introduced when measur¬ing the cross-sectional area of the airway or entry. Under the best of circumstances, measurement errors, instrument errors, and a host of other minor errors all combine to cause a total error of at least 10% in the velocity calculation. The vane anemometer also can be used with reason¬able accuracy to measure airflows in mine-ventilation ducts. In this application, the anemometer is mounted on a rod and is held at the center of the duct end. For a duct that is discharging air, the average velocity in the duct is 85% of the centerline reading (Northover, 1957). For a duct that is taking in air, the average velocity is 70% of the centerline reading (Haney and Hlavsa, 1976). To measure the airflow discharged from a regulator or from a small hole in a stopping or bulk¬head, a correction factor for the area is necessary. A good approach in this situation is to traverse the area of the regulator or hole, holding the anemometer with an extension rod. This provides an average velocity that is multiplied by 85% of the measured area of the regulator or hole. In all cases, the manufacturer's instrument cor¬rection table must be used and applied properly. For accurate results, the anemometer should be returned to the manufacturer for periodic cleaning and checking. If it is in daily use, the anemometer should be returned about once per year, and proportionally less frequently if the usage is less frequent than on a daily basis. Smoke Tube The smoke tube may not appeal to individuals who believe that good measurement results can be obtained only with expensive, complicated, and fragile instru¬mentation. Nevertheless, smoke works about as well as anything for the routine measurement of low air velocities in mines. The following procedure yields reasonably good results: 1) Two marks are scratched 7.6 m (25 ft) apart on the floor of the airway. 2) The smoke tube is used to release a cloud of smoke in the center of the airway, about 0.9 m (3 ft) upstream of the first mark on the floor. 3) A timed interval begins when the leading edge of the smoke cloud passes over the first mark, and the interval stops when the leading edge of the cloud passes over the second mark. 4) A factor of 20% is subtracted from the cal¬culated velocity to determine the true average velocity, providing a correction for the centerline and for the spreading effect at the front of the cloud. Velocities calculated with the preceding method generally are accurate to within 10 to 15%. In some instances, the cloud from a conventional smoke tube

Jan 1, 1982

By B. Anameric, S. K. Kawatra

Pig iron nuggets were produced in a laboratory scale furnace at Michigan Technological University. These nuggets are produced from pellets consisting of a mixture of iron oxide, coal, flux, and a binder, which are heated to 1450°C.These pellets then self-reduce to produce a solid, high-density, highly metallized (96.5% Fe) pig iron. During the nugget production process, a separate slag phase forms that cleanly separates from the metal. The physical and chemical properties of the pig iron nuggets are similar to pig iron produced by blast furnaces, which is distinct from Direct Reduced Iron (DRI).

Jan 1, 2004

By T. R. von Ow

The method of powering mills such as AG, SAG and Ball has a long and technically interesting history. It can be said that almost every decade in the last half century had its own trend. The trends were dominated by the market demand, the advanced process development and requirement of the mechanical and the electrical technology but mostly by the price of the applied technology. The increasing size of the ring geared mills in the last years requires more awareness on the drive train solution. Special attention must be paid to the mechanics and the requirements of the power grid. Operational challenges such as soft starting and controlled operation as well as the protection of the equipment have gained importance. Frequency converter drive solutions dedicated to single or dual pinion ring-geared mills offer significant advantages over conventional drives due to their application specific features and controls. Solutions with low speed motors as well as high speed motors cover all possible configurations available for ring geared mills.

Jan 1, 2010

By Jim Oehler

The US minerals industry, as well we know, has been hit hard by the recent recession and is faced with what is projected to be a slow recovery. Even so, there was an underlying feeling of optimism at American Mining Congress' International Mining Show in Las Vegas, NV, Oct 11-14. More than 25,000 people attended during a particularly rough financial year, dispelling fears that attendance would dip sharply below 1978's 30,000 plus tally. The equipment expositionheld every four years-was the largest show of its kind this year in the Western Hemisphere in terms of exhibit space, providing an extensive view of mining-related equipment. In addition to the nearly 700 exhibits occupying more than 445 hm2 (11 acres), 20 technical sessions were held covering topics ranging from minerals availability and the state of the mining industries to communications, computers, and accounting. Government officials were well represented, with six US senators and five US representatives participating in various convention functions. Sen. Paul Laxalt (R-NV) was the keynote speaker at the welcoming luncheon. Other senators involved in the show included Howard W. Cannon (D-NV), Pete V. Domenici (R-NM), James A. McClure (R-ID), Frank H. Murkowski (R-AK), and Steven D. Symms (R-ID). Representatives on the program were Joseph M. Gaydos (D-PA), Guy Vander Jagt (R-MI), Don Young (R-AK), Jim Santini (D-NV), and Ceil Heftel (D-HI). There were also many international vistors at the show, including exhibitors from Australia, Austria, Canada, England, Finland, Sweden, South Africa, and West Germany. Equipment Highlights Equipment and services featured at the exposition ranged from aerial mapping to wire rope. Aisle space alone between exhibits exceeded 8 km (5 miles). And the companies represented ranged from small businesses to large corporations. Some of the company products described here are representative of the many exhibitors in attendance. Atlas Copco's extensive display of rock drilling and tunnel and shaft boring equipment featured its Boomer H-201, the newest addition to its line of underground drifting rigs; the ROC 306, a new track-mounted, high pressure down-the-hole drill rig; a new hydraulic surface crawler drill, the ROC 712H, designed for small and medium-size holes in quarries and small open-pit mines; a computerized analysis system, CARE (see page 1570), designed to project productivity and operating costs for various mining operations; several electro-hydraulic loaders, including the new rail bound LM E60 and LM E80, and two Hagg-loaders-the 7HR and 10HR-that use two digging arms to rake and guide muck; a mechanized blasthole charging system; rock bolting equipment and its new Swellex roof bolting method; Sandvik Coromant drill steel and bits; and its new Jarva R200 raise boring machine.

Jan 11, 1982

By Menglong Bian, Xiaotong Dong

Thermal damage is an urgent problem faced by deep mining. High-temperature rock mass, hot water and electromechanical equipment are the sources of the thermal damage. The increase in air temperature in mine tunnel is the result of fluid–solid coupling heat transfer from various heat sources. This paper takes a roadway with hot water and a mine car as the research object. Based on the Fluent, Static Structural and Steady-State Thermal modules in ANSYS, the thermal-fluid–structure coupling analysis in the roadway is carried out. It is found that the hot water has a significant effect on the thermal environment of the mine. Under the experimental parameters, the maximum heat flux density and equivalent elastic strain at the interface between the hot water and the rock are 9789.3 W/m2 and 0.00054532 m/m, respectively. The heat flux distribution in the rock area near the mine car does not change much, and the equivalent elastic strain value in this area is high. After the air is heated in the hot water area, the temperature curve changes gradually and then rises steadily. By changing the inlet air temperature parameter, it is found that the lower the inlet air temperature, the better the cooling effect, and the change of the inlet air temperature can change the air temperature field in the roadway as a whole. By changing the parameter of inlet air speed, it is found that higher wind speed brings better cooling effect, but with the increase in wind speed, the cooling effect of roadway thermal environment is very limited. When the wind speed increases from 5 m/s to 10 m/s, the air temperature of most measuring points drops by more than 2℃; when the wind speed increases from 15 m/s to 20 m/s, the air temperature of most measuring points drops below 0.7℃. After the wind speed reaches a certain value, the cooling effect of ventilation is very limited.

May 5, 2022

By Michael A. Linkous, Mark J. Zdunczyk

Silica in the form of sand and sandstone is one of the most common, and at the same time, unique industrial minerals. Found in every rock type of every geologic age and virtually everywhere in the world, silica is used in products that touch just about every aspect of daily life. Imagine a world in the 1990s without computer chips, fiber optics, or glass, and you have just begun to understand how important silica is to the quality of life we enjoy. The elements silicon (Si) and oxygen (0) comprise roughly 60% of the earth's lithosphere to a depth of about 16 km. The crystal structure of silicon dioxide consists of one atom of silicon bonded to four surrounding atoms of oxygen to form a three-dimensional network of SiO, tetrahedra. This network makes up the mineral quartz (Murphy and Henderson, 1983), the most common detrital mineral in most sandstones. Quartz is also a major constituent of many igneous and metamorphic rocks and is widespread as a siliceous cementing agent in various rock types. Although quartz is common, sandstones, quartzites, and pegmatites and the unconsolidated sediments derived from them that have a silica content high enough and pure enough to meet today's market demands for quality and consistency are not common. USES AND SPECIFICATIONS Silica sand that is mined and processed for industrial uses must conform to the chemical and physical specifications set by customers. In the United States almost half of the silica sand produced is used in the manufacture of glass. Other important products include foundry sand, ground silica, blasting sand, and fracturing sand. Glass Sand Silica is the principal glass-forming oxide in a glass batch. Glass manufacturers develop model specifications for each source of silica sand used. These specifications broadly define the limits and ranges for chemical and physical properties of the sand and are used by the manufacturer in calculating the desired batch mix or formula. Some specifications may be critical to a glassmaker and require very stringent limits on the quantity of impurities in the sand. For example, the total iron oxide content of a batch is extremely crucial when making white or flint glasses (Mills, 1983). Iron is present in almost every raw material used in a glass batch and must be carefully controlled in order to obtain a consistent color in the finished product. It is difficult, however, for a raw material supplier to tightly control the chemistry of a naturally occurring material such as silica sand. To a great extent the commercial quality of a sand is determined by its geologic history. Realizing this, glass producers tailor their model specifications to each source of approved material. In general, a glass company is concerned most about the consistency of raw materials on a day-to-day basis. Soda-lime-silica glass was the earliest type of manmade glass (Baker-Can, 1967) and still accounts for most of the glass manufactured for commercial use today (Mills, 1983). It is relatively easy to melt and shape and is less expensive per ton to produce than most other types of glass (Baker-Can, 1967). Soda-lime-silica glass is used in fabricating containers, flat glass products, incandescent and fluorescent lamps, glass fiber, and many other products. Heavy minerals such as ilmenite, leucoxene, kyanite, and zircon are impurities on which strict limits are placed for a glass batch. Because of their refractory nature they either do not melt or only partially melt, which results in stones or feathers in finished glass. Aluminosilicates such as kyanite also contribute unwanted alumina to the batch as they partially melt. Limits are especially rigid for refractory mineral grains coarser than 0.60 to 0.425 mm (30 to 40 mesh). [Tables 1 and 2] present typical specifications for silica sand used in flat glass and container glass products. The percentages shown represent an average of many companies7 specifications.

Jan 1, 1994

By Lawrence M. Brennan

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

Jan 1, 1976

By Rodrigo Peroni, Euler Victoria

Knowing the depth of contacts between weathering zones in phosphate mines is determinant to define the stripping ratios and the mineralized zones. Contacts are gradational and difficult to define even in the drillhole samples, so the combination with geostatistical simulation techniques helps to define the contact zone and the grade distribution, providing a way to assess the probability of the occurrence and the positioning of contacts between the barren waste and the ore zones. These inputs are used in the scheduling, anticipating risks associated with mine planning scenarios. The results from an application to a case study showed a good improvement when compared to traditional techniques.

Jun 25, 2023

By Thornton L. Neathery

A small area in the Alabama Piedmont was studied to determine if metallic enrichment was associated with local magnetic anomalies. Cores from two holes in these magnetic anomalous areas have sulfide (pyrite, pyrrhotite, chalcopyrite) film on slicken-side and fracture surfaces. Two hundred and fifty soil samples were collected from the top of the "B" horizon on 1, 000 foot centers over the general trend of magnetic anomalies. Soil samples were leached with a 1:2. 7 mixture of HF and H2SO4 and the dilute filtrate analyzed for Cu, Pb, Zn, Ni, Co, Cr, Mo, and Cd, using atomic absorption methods. Concentration-value maps were prepared for each element. Small areas of anomalous concentrations were defined for Cu, Pb, Zn, Mo, and Ni. Some of these areas generally coincides with magnetic anomalies. Concentration values (ppm) range from: Cu 165.0 - 0. 5; Pb 265. 0 - 0.0; Zn 1386.0 - 43. 0; Ni 187.0 - 10.0; Co 207.0 - 6. 5; Cr 117.5 - 2.0; Mo 95.0 - 0.0; and Cd 32. 5 - 2.0. Studies of other magnetic anomalies appear warranted.

Jan 1, 1972

By Elza Burch

Introduction The feeding to zero concept involves feeding 600 µm x 0 (28 mesh x 0) size raw coal to heavy media (magnetite) cyclones along with the +600 µm (+28 mesh) size coal. Traditional circuits employ desliming or removing the 600 µm x 0 (28 mesh x 0) size fraction and feeding the cyclones +600 µm (+28 mesh) size coal. The feeding to zero concept recirculates 600 µm x 0 (28 mesh x 0) fines in the circuit. At the same time, a portion of the fine material is continuously withdrawn and recovered. This, in turn, prevents a fines buildup. This concept eliminates desliming screens and secondary fines circuitry for recovery of 600 x 150 µm (28 mesh x 100 mesh) coal. The result is a very simple circuit. Feeding to zero at Island Creek Island Creek Corp. was the first involved with the new concept in 1976. The company needed a temporary plant for the 9.5 mm x 0 (0.4 in. x 0) raw coal at its Pond Fork mine, near Madison, WV, while a full-scale plant was being designed and built. At that time, the Childress Corp., of Beckley, WV, became interested in the feeding to zero concept. Island Creek awarded a contract to Childress to build a single cyclone modular plant, incorporating this feeding to zero concept. The plant was erected in three months. The Pond Fork modular plant proved successful in attaining the desired feed rate of about 63.5 t/h (70 stph), while maintaining good separating efficiencies and low magnetite consumption rates. The 9.5 mm x 150 µm (0.4 in. x 100 mesh) clean coal was recovered and the 150 µm x 0 (100 mesh x 0) size was disposed of to waste. The full-scale plant was completed about two years later and the Pond Fork modular plant was moved to Holden, WV. There, it was incorporated into the Holden 29 preparation plant as a separate circuit for cleaning -25 mm (-1 in.) coal. In 1976, a similar plant was installed in Virginia by another company. These two plants are believed to be the first two operational plants in the United States incorporating the feeding to zero concept. Island Creek subsequently contracted with Childress for an identical plant at the Coal Mountain operation in West Virginia. The plant operated for three years before the mine was closed. The unit was then moved to the Spurlock mine near Martin, KY where it continues to operate. The successful operation of the Pond Fork and Coal Mountain plants before and after relocation proved both the performance and moveability of this type of circuit when constructed in a modular fashion. Since the first two plants were built, Island Creek has incorporated the feeding to zero circuit in nine additional plants. A grand total of 33 cyclones have been installed using this concept. One is the Providence mine, near Providence. Providence preparation plant Island Creek contracted with J.O. Lively Corp. of Glen White, WV in July 1978 for the construction of the Providence preparation plant. The plant began operation in February 1979. The construction period was about halved by building the plant with modular design concepts. Prefabricated sections, floors, and sides were brought in as units and then bolted in place. The Providence plant has a good track record of processing coal at a feed rate of 454 kt/h (500 stph). Feed coal to the plant has an average ash content of about 18% and sulfur content of about 4.5%. It contains about 22% refuse. The coal product has an average ash content of about 8% and sulfur content is about 3%. Raw West Kentucky No. 9 seam coal is conveyed from a box cut in the Providence mine to a rotary breaker. The breaker is fitted with 74 mm-diam (3 in.-diam) opening breaker plates. Therefore, it is well suited for removing trash, roof bolts, wood, and pyritic balls that are common in Illinois Basin coal. The -75 mm (-3 in.) coal is conveyed from the rotary breaker to

Jan 8, 1987

By Mark Van Dyke, Ted Klemetti, Jake Beale, Zoheir Khademian, Joe Wickline

In 2016, a 3.7-ML magnitude event caused by mining activity occurred at a longwall mine in southwestern Virginia which was recorded by the United States Geological Survey (USGS) and felt by local residents. The event was the largest of its kind since a global mine design change by the operator was instituted in 2008 following three large events in 2005, 2006, and 2007 (3.4, 4.3, and 3.4 Moment Magnitude (Mw), respectively). Two of the three pre-2008 events (2005 and 2007) damaged ventilation controls in the mine which fueled a mine fire. In 2016, the mine’s management requested researchers from the National Institute for Occupational Safety and Health (NIOSH) to access geological data and determine what parameters could possibly lead to events of a magnitude of 1.0 ML or greater. Evaluation of 2152 geological data points and modeling revealed three major geological factors in common with the majority of the 181 recorded +1.0 ML events from 2009 through 2016. Three levels of seismic potential were identified as follows:

Aug 4, 2023

By J. M. Mutmansky

Introduction Since the Hardy Cross numerical algorithm was applied to the solution of mine ventilation networks, numerous developments have taken place in ventilation network analysis procedures. The Newton-Raphson method, the linear theory method and the second-order approximation method are three of the additional approaches that have been applied. The efficiency of a ventilation network analysis procedure depends not only on the method of solving the system of equations derived from ventilation networks but also on the method of deriving the equations through selection of the meshes. This paper investigates the impacts of three mesh selection methods on the efficiencies of four ventilation network procedures. The minimum-resistance spanning tree method is the most popular procedure for selecting meshes. The shortest-path method was suggested by Epp and Fowler (1970) for improvement of the Newton-Raphson method. The minimum-resistance-path method is suggested here as the third method of selecting meshes. The effects of these three methods on the Hardy Cross method, the Newton-Raphson method, the linear theory method and the second-order approximation method are analyzed in this paper. These network analysis methods have been summarized in the paper by Kim and Mutmansky (1991). Mesh selection methods A procedure for mine ventilation network solution based on mesh equations typically consists of two steps: • the mesh selection step and • the network solution step. The mesh selection step establishes a set of independent equations to solve the network. The resulting equations are nonlinear. One method of solving nonlinear equations is to linearize the equations and iteratively solve these linear equations. The four analysis methods previously discussed in this paper use this approach. The linear equations required for solution can be written in matrix form as: P•dQ = dF (1) where P is the pseudo-resistance matrix, Q is the quantity matrix, and F is the Jacobian matrix (Kim and Mutmansky, 1991). The efficiency of the solution method depends on the characteristics of the P matrix, which in turn depend on the method of selecting meshes. To understand the impacts of the mesh selection methods on the efficiency of the ventilation network analysis procedures, the characteristics of the P matrix must be understood first. The diagonal elements of the P matrix, m in number, are the sums of the pseudo-resistances of the branches contained in the m meshes necessary for solution. The off-diagonal elements are the sums of the pseudo-resistances of the branches shared by two meshes. For example, P12 (first row, second column element of the P matrix) is the sum of the pseudo-resistances of the branches common to mesh 1 and mesh 2. When solving the system of equations expressed in Eq. (1), three of the iterative procedures (the Newton-Raphson, the linear theory and the second-order approximation methods) use the direct factorization method (Burden and Faires, 1985) on the P matrix. When the direct factorization method is used, the computational efficiency is enhanced by larger diagonal elements and a sparser P matrix. Generally, the sparsity of the P matrix is more important than the size of the diagonal elements. There is no known mathematical method of predicting the efficiency of these procedures, but heuristic methods can be used to study the efficiency. The Hardy Cross algorithm applies the Gauss-Seidel method of splitting the P matrix and applies the Newton-Raphson procedure. The efficiency of the Hardy Cross method depends on the spectral radius of the matrix[[D-L]-1U], where D is the diagonal, L is the lower triangle and U is the upper triangle of the P matrix for the Newton-Raphson method (Ortega, 1972). The smaller the value of the spectral radius, the faster the convergence. As with the other three methods, larger diagonal elements and a sparser matrix aid in faster convergence. With the Hardy Cross method, however, the size of the diagonal elements is more important than the sparsity of the matrix. Remembering that each of the diagonal elements of the P matrix is the sum of the pseudo-resistances of the branches in a mesh and that each off-diagonal element is the sum of the pseudo-resistances of the branches shared by two different meshes, the following strategies can be used: • Avoid having branches with larger resistance values shared by multiple meshes. This increases the size of the diagonal elements and reduces the chance of off-diagonal elements being large.

Jan 1, 1993

By Günther M. Volkmann

"INTRODUCTION The construction of the BAB 44 motorway between Kassel and Herleshausen (both Germany) is an important part of the trans-European road network. After its completion, the A 44 will permit an efficient connection of the Benelux States in the West and Poland in the East. In the section between Eschenstruth and Hessisch Lichtenau the Hirschhagen Tunnel is constructed with a length of 4.1km. Most sections of the tunnel are situated in weak ground conditions. Hence a combination of face bolts with spiles (short forepoling) or pipe umbrellas (long forepoling) additionally support the tunnel heading on a regular basis This article will focus on the application of a newly developed method to connect the pipe umbrella support, which is mostly installed by conventional drilling machines nowadays. The installation method limits the outer diameter to a maximum of 168mm and a pipe wall thickness of 12.5mm because workers must be able to handle pipes and its drill steel. Typically these pipes are connected with threads that create a weak link in the support system so during the last decade alternative connection systems were developed to overcome this disadvantage. One of these alternative coupling systems is the so called Squeezed Connection, which will be explained in this article followed by a discussion of its advantages and disadvantages regarding technical specifications, construction time, and safety issues. THE HIRSCHHAGEN HIGHWAY TUNNEL The German highway A 44 directs from Aachen through the Ruhrgebiet to Kassel (all Germany) and it is extended to Herleshausen (Germany). The new section connects the A 7 to the A 4 after completion. This upgrades the east-west highway connection starting at the Benelux Union and heading to Poland so it is an important part of the trans-European road network. The newly constructed section has a total length of approximately 63 km and comprises 13 tunnels. All tunnels are constructed as double tube tunnels with 2 lanes and emergency sidewalks on both sides and 14 emergency cross passages. The Tunnel Hirschhagen starts in the north of Eschenstruth and ends near Hessisch Lichtenau (Figure 1). The length of the northern tube is 4,147 m and that one of the southern tube is 4,204 m. So it is the longest tunnel of the A 44 and the second longest highway tunnel in Germany when it is put into operation. On its alignment it crosses underground the Losse valley with the state road 7, the streetcar Kassel-Hessisch Lichtenau and the Losse stream. The cover above crown varies from 10 m to 95m and it is the lowest when crossing the Losse stream."

Jan 1, 2016

By Edward M. Dowey

This paper presents the excavation of the Roosevelt Island Valve Chamber, the final of three chambers in Stage One of New York City's Third Water Tunnel. The valve chamber and passageways 13 379.7 cu m (17 500 cu yd) are drilled and blasted. The rock is stabilized by grouted rock anchors. The 6.1 m (20 ft) diameter shaft is drilled, blasted, rock bolted and wire meshed for a depth of 70.7 m (232 ft). Then shaft excavation proceeds with immediate concrete lining for another 120.7 m (396 ft). Two 6.1m (20 ft) diameter manifolds are excavated using heading and bench. The raise drilling of eight 3.7 m (12 ft) diameter laterals occurs during lower shaft sinking.

Jan 1, 1987

The Colorado portion of the Rio Grande rift is clearly defined physiographically from the south in the San Luis valley northward to the latitude of Leadville where sediments of the Dry Union Formation (Tweto, 1961) accumulated within the subsiding rift zone. To the north of Leadville is a system of faults active in the late Tertiary, including the Mosquito fault that offset a portion of the Climax mineralization (Wallace et al., 1968). Tweto (197B) describes remnants of Dry Union Formation against the Mosquito fault north of Dillon in the Blue River valley, so it is apparent that the rift and associated sediments extended well to the north of Leadville. Kelley (1952, 1954, 1956) and Chapin (1971) have noted in New Mexico that the Rio Grande rift is a series of en echelon basins, offset to the east, with constrictions connecting them (Fig. 1). In Colorado the structural basins are offset to the west. The net effect of all the rift zone basins in New Mexico and Colorado is an arcuate pattern (Fig. 1) abutting the Colorado Plateau and Sari Juan Volcanic Province on the west. The width of the rift system gradually decreases northward to about Leadville where rift definition is not apparent.

Jan 1, 1987

By L. R. Gremillion

Chemical and mineralogical data on 25 apatitic phosphate ores of diverse origins from commercial deposits around the world were widely variable. From 14 to 18 chemical constituents of each sample were determined, and each sample contained 4 to 10 mineral phases. Altogether there were 34 different mineral phases, many of which occurred in only one sample. The samples contained varieties of francolites, fluorapatites, and mixed hydroxyl-fluorapatites. The apatite compositions calculated from the chemical data showed that certain types of substitutions were characteristic of different apatite varieties, but every variety had some degree of substitution of (CO3)2 for (PO4)3 and the P2O5 contents of the apatites ranged from 35.1% for a North Carolina apatite to 41.9% for a Missouri apatite. In a concentrate with a grade of 32% P2O5, the North Carolina sample should contain about 8.7% gangue and the Missouri sample about 23.7% gangue. To optimize the digestion process of a phosphate are, it is necessary to consider the dissolution behavior of the different mineral phases during acidulation and the effects they may have on the chemical properties of a fertilizer intermediate. In the production of wet-process acids or superphosphates, the acidulation of certain minerals may cause foaming and formation of sludge, filtration, corrosion, and handling problems.

Jan 1, 1978