Search Documents

By George H. K. Schenck

INTRODUCTION Boom-type miners mechanize the cutting and loading of a solid face of rock having a compressive strength as high as 124 MPa (18,000 psi). These machines differ from continuous miners in that they can cut much harder rock for any given cutter-motor power. In addi¬tion, a boom miner is more flexible, selectively cutting any cross-sectional configuration from circular to square arched faces, 1.57-rad (90°) turns, and mixed faces. The basic operational concept of these machines, utilizing a rotary cutterhead mounted on a slewing boom, gained importance in Europe during the 1960s, where, due to their specialized use in roadway development in coal mines, they were called roadheaders. Along with continuous miners, boom miners share the ability to fully mechanize the cutting and loading of rock, while eliminating the need for explosives and the shattering and overbreaking of roof and rib rock that is normal with conventional mining methods. The machines are crawler-mounted and most have a gathering-arm load¬ing system that, moves the cut rock onto an internal chain conveyor, subsequently discharging the rock into a transport device such as a shuttle car or a bridge conveyor. Light- to medium-weight boom miners are cutting softer rocks and minerals successfully, including coal, shale, rock salt, potash, phosphate, weak sandstones (e.g., those associated with uranium ore), and other softer ores such as iron. Heavy third-generation ma¬chines have extended rock cutting to much harder ores such as borate, oil shale, and almost all rock associated with coal seams, except for the harder sandstones and limestones. The mention herein of specific manufacturers, brand names, or models is made by way of example only and does not constitute endorsement by the author or The Pennsylvania State University. All costs cited herein are in 1977 US dollars. TYPES OF BOOM MINERS All boom miners utilize drag bits fixed on a rotary cutterhead that is powered by an electric motor mounted in the slewing and elevating cutter boom. The cutter¬head is either a ripping or a milling type, and the machines have evolved through three generations (Boldt, 1977). As shown in Fig. 1, only one boom miner cur¬rently produced can use both ripping and milling cutter¬heads on one machine. Table 1 lists boom-miner models built in large num¬bers and/or successfully used or marketed in North America. Table I excludes machines of which only a few have been produced and none sold in North America, such as those manufactured by Demag, Eick¬hoff, Paurat, and Thyssen-Rheinstahl. The only Ameri¬can boom miners are the F-16 and Super ROC¬MINER"', manufactured by AEC, Inc., State College, PA. Machine Generations Three generations of boom miners can be distin¬guished according to their times of introduction, their rock-cutting abilities, and their weight classifications. Fig. 2 illustrates typical best-practice production rates possible from boom miners for selected cutter-motor power sizes (other power figures can be interpolated). Productivity is related directly to the energy applied to the cutterhead and indirectly to machine weight; higher

Jan 1, 1982

By Jerzy Iskra, Wlodzimierz Sikora

INTRODUCTION The effect of stress concentration in coal seams leads to a sudden release in the form of rock bursts. This effect is controlled along many lines, but the most efficient methods consist in preventing the accumulation of energy in the seam or in the roof. The over- burden weight resulting from mining is the main cause of excessive stresses in a seam. An additional weighting of roof takes place due to formation of abutment loads. The accumulation of energy in the roof takes place where thick sandstone beds occur in overburden strata. An essential problem is to watch, with appropriate measurements, the stress of coal and roof rocks, and to conclude, on this basis, the degree of rock burst hazard. The methods for determining the stresses in a coal seam are quite well developed at present. Let me mention here the seismoacoustic methods, or borehole overcoring measurement methods. On the other hand, an examination of the state of stress in the roof rocks is almost impossible at the actual stage of development of the measuring technique. Hence, the model of stresses in roof was analyzed and determined theoretically. Based on this model, a technology for roof fracturing was developed which aims at destressing rock masses in a planned way. The essentials of the fracturing method consist of a controlled breaking of sandstone roof beds in order to: - prevent the excessive weight on the pillar - prevent a spontaneous breaking of roof beds at a limiting state followed by sudden energy release. High pressure fluid infusion is used to release buildup of stresses. THEORETICAL AND RESEARCH BASES According to the tests carried out by A. Kidybinski by the method of finite elements, isobars of stresses occurring in the roof over the working face are shown in [Fig. 1]. The results indicate: - the highest values of compressive stresses occur in the roof directly over the seam, at a distance of about 5 m ahead of the working face - at a distance of about 22 m above the seam there is a line of transition of compressive stresses into tensile - within the tested roof area, the highest tensile stresses occur at a vertical distance of 35 m, and at a horizontal distance of 20 m from the line of the working face. H. Gil (1) determined the pres- sure initiating the fracture of roof rocks by the formula: [ ] On the basis of results of analysis of isobars, and of the initiation of failure stress formula (1), isobars may be designed for a given roof class. Assuming for sandstones the Poisson ratio equal 0.3, a field of failure stresses presented in [Fig. 2] is obtained. In this figure it can be seen in which areas the effects of sandstone breakage may be obtained with the operation of a determined hydraulic pressure. To use formula (I), the knowledge of Rr is still necessary. This value may be determined in two ways: - taking a sample and testing it in the laboratory, if there is access to roof rocks in the existing openings, - drilling boreholes in the roof from surface and examination of the rock mass by using hydraulic breaking, if there is no access to roof rocks.

Jan 1, 1982

By Frank W. Schmechel

The subsidence monitoring program at the Old Ben No. 24 Mine was planned in 1976 as part of the on- going demonstration of longwall mining. The construction and installation of the network of surface subsidence monuments was started in Play of 1976 and completed four weeks later after which an initial reference survey was made of all surface monuments. Mining of the first longwall panel started in September of 1976 and was completed in May of 1977. Mining of the second panel started in August of 1977 and is continuing at this time. The Automatic Data Acquisition System (ADAS), which is the subject of this paper, was installed over the second panel on March 29, 1978 near the panel center (Figure 1) and approximately 104 meters (370 feet) in front of the face position at that time. Mine No. 24 is located approximately 5 kilometers (3 miles) Northwest of Benton, Franklin County, Illinois. Coal is being mined from the Illinois (Herrin) No. 6 seam which is at a depth of about 189 meters (620 feet) at the site. The Herrin No. 6 is up to 2.7 meters (9 feet) thick in the mine but a constant thickness of 2.1 meters (7 feet) is being mined by the demonstration longwall with both roof and floor coal being left for stability. Although the seam is flat laying at the mine, it dips north- ward regionally. Topographically, the surface over the demonstration panels is essentially flat and low laying with several areas of standing water. The Rend Lake Dam (United States Army Corps of Engineers) is about 305 meters (1000 feet) north of the demonstration site.

Jan 1, 1979

By Brett P. Spigarelli

To meet the growing need for CO2 capture and storage technology, Michigan Technological University is researching CO2 capture and storage using alkali solutions1. Without implementing emission controls, world energy emissions will rise from 29.7 billion t in 2007 to 42.4 billion t in 2035. The first objective of the project was to determine the optimal amount of alkali in solution for CO2 absorption. This was accomplished using theoretical equilibrium calculations and experimental absorption data for alkali solutions of 1, 2, 3, 4 and 5 weight %. As the % alkali in solution increased, so, too, did the ionic strength of the solution. Theoretical calculations took into account the effect increasing ionic strength will have on the solubility of gaseous CO2. The calculations assumed negligible formation of H2CO3, constant temperature and constant pressure. In the experimental setup, a 16% CO2 /balance air gas mixture was bubbled through an Erlenmeyer flask filled with an alkali solution. The gas exiting the flask was fed to a gas analyzer to record the % CO2 in the exhaust gas stream. The % CO2 absorbed was taken as the % CO2 in the exhaust gas minus the % CO2 in the feed. Theoretical calculations found that as the % alkali in solution increases, the solubility of CO2 decreases. This was verified by experimental data, which found that a 2% alkali solution, by weight, was the ideal solution for CO2 absorption. Current research is focused on using the 2% alkali solution for permanent storage of CO2 via mineral carbonation to produce a value-added product.

May 1, 2012

By A. Frederick Banfield

Various modeling techniques are available for describing ore deposits on the computer. Multiple seam coal deposits can be quite successfully modeled by using the variable block modeling method. In essence, this type of model provides an efficient method for storing the relevant data required for calculating ore reserves and for performing the mine planning. The main advantages of the variable block model over other methods for modeling irregular block or tabular deposits are: 1) The amount of data that has to be stored is minimal, i.e., the most efficient use of computer space is made since only the pertinent seam and topographic geometries need to be stored. 2) The computer logic necessary for calculating ore reserves and performing mine sequencing is relatively simple as compared to the logic that would be required for the three-dimensional block model. 3) A variable block model can be created either by starting from raw drill-hole data or by digitizing cross sections that contain the geologic interpretations. 4) Updating individual or multiple blocks can be performed with relative ease. 5) Display of the model is relatively inexpensive and easy to perform. The objective of this chapter is to describe a developed procedure as used in applying the variable block model to a specific multiple seam coal deposit. The general sequence of operations includes: (1) develop the computer model, (2) design the open-pit limits, (3) compute reserves, and (4) develop a long-range mining plan. The programs developed for this application are an expansion of MINTEC's MEDSYSTEM software. It should be noted that this is just one of many possible applications of the developed variable block model.

Jan 1, 1979

By N. S. Tanious

Three water-spray ventilation systems were tested to determine their effectiveness in diluting methane at a room and pillar coal mine face. These systems were: the shrouded nozzle, the spray fan, and the conventional nozzle system that is normally supplied with continuous miners. The testing was carried out in two stages. The air-moving capacities of the nozzles were first evaluated in a duct. The duct tests showed that: (1) the volume of air moved by the nozzle is controlled by the type of nozzle, the space around it, and its location relative to the duct wall; (2) the spray fan nozzles or the shrouded nozzles are both capable of moving air quantities up to 6400 cfm (3.02 m3/s); and (3) when the shroud was removed from around the shrouded nozzle, its bare nozzle moved between 9 and 65 percent more air. In the second stage of testing, a full-scale evaluation of the three systems was conducted. The three spray systems were mounted on a continuous miner which was positioned in an aboveground simulated mine entry. The simulated face tests gave the following results: (1) the standard nozzle system, which is mounted on the front end of the miner, is inadequate for diluting methane; (2) under the best testing conditions (lowest methane emission rate, highest primary ventilation airflow, and closest brattice set back), the slight improvement given by either the shrouded system or the spray fan is the same; and (3) under the worst testing conditions (highest methane emission rate, least primary ventilation airflow, and farthest brattice set back), the spray fan gave superior performance.

Jan 1, 1987

The 70th Annual Meeting of the Minnesota Section of SME and the 58th Annual University of Minnesota Mining Symposium were held April 22-24, 1997, in Duluth, MN. About 450 attended both. The Fifth Annual Trade Expo, held during the meeting, was attended by a record 80 exhibitors. An SME short course on materials handling preceded the meeting. Iron-mining and the steel industries highlighted the meetings. Specific topics in the keynote addresses and technical sessions primarily concerned iron mining, iron- ore processing, product (taconite pellet) transportation and the steel market. Although the iron mining industry held center stage, several papers outside of iron mining were also presented. Iron mining in Minnesota contributes $1 billion to Minnesota's economy. According to the Iron Mining Association of Minnesota, more than 20,000 state jobs rely on the industry. Minnesota's iron mining industry came through tough times in the early 1980s, mostly because of worldwide competition. However, the industry has shown significant improvement in the last 15 years. A large share of the steel industry's competitiveness in the late 1990s is due to cost and quality improvements in the taconite mines. The taconite industry has become more productive, largely through the application of new technology. Since the early 1980s, Minnesota's taconite mines have nearly doubled the number of tons per worker-hour. Most of the taconite reserves in the United States are located in Minnesota's Mesabi, Vermilion and Cuyuna iron ranges. The Mesabi range alone contains an estimated 24. Gt (23.6 billion long tons) of mag- netic taconite. Since the 1950s, when taconite production first began, 1 Gt

Jan 1, 1997

By R. P. Benson, D. H. MacDonald, D. R. McCreath

1 - INTRODUCTION The Canadian Shield is well known as being among the older exposed areas of the earth's surface. Except for local zones, its extent from the Arctic to central North America is characterized by strong and durable crystalline rocks. It is generally regarded as highly competent for tunnelling and underground excavation and, as a result, many openings are excavated in it with minimal temporary support. If hydraulic efficiency is not a factor, tunnels do not usually require continuous concrete linings. A high percentage of the crystalline rocks of the Shield are of metamorphic origin and as such one of the dominant characteristics is their foliation. Zones of sheared or faulted rock are common in the Canadian Shield, particularly in association with foliation. These zones may be created during the late stages of the metamorphic process, leaving shear zones as an integral feature of foliated rock. As foliation is a preferred direction of weakness, deformation and rupture during subsequent tectonism may also create parallel shear zones. These "foliation shear zones", as described by Deere (1973), are an adverse feature of the otherwise durable rocks of the Canadian Shield. The zones may range from faults of major proportions to seams only a fraction of an inch thick. If fault gouge or alteration products exist in the zone the strength of the rock mass may be great1y reduced. As foliation is usually continuous, large unstable wedges of rock may be created if the opening is subparallel to the foliation. The effect on the construction of underground openings in terms of excavations, support and scheduling can be significant.

Jan 1, 1974

By Robert C. Jones

The Henderson Mine, a new large scale underground development, is being developed by AMAX. The mine site is located about 50 miles west of Denver, Colorado, adjacent to the east side of the Continental Divide. The mine's three access shafts are in a narrow alpine canyon at an elevation of 10,400 feet. The ore body of molybdenite is to the south of the shafts, 3,000 feet below the surface of nearby Red Mountain. The ore extends over 800 feet vertically and is about 1,700 feet by 2,200 feet horizontally. The eastern and southeastern boundaries of the deposit are still to be defined at the present time. Development mining is advancing at the mine site on a three shift, seven-day basis at the rate of about 2,000 feet per month. About twelve miles of heading have been driven at the present time, and work to sink No. 3 Shaft has started. Development mining was begun in 1967 with the sinking of No. 1 Shaft. This shaft is fully concreted, 23 feet in diameter and 2,441 feet in depth. It is the mine ventilation exhaust conduit. It is also used for development muck hoisting, stand-by mine water pumping and secondary access to the 8,035 foot level. This level is the mine exhaust ventilation level and was mined principally by conventional rail mining methods. From this one station at the 8,035 foot elevation a 5% ramp was driven to the initial production horizon at 8,100 feet with diesel rubber tired mining equipment. Subsequent mining has been entirely by mobile diesel equipment at the mine site.

Jan 1, 1974

By Adrian C. Dorenfeld, John T. Oleksy

John D. Maefadyen Subject paper unfortunately has some serious flaws negating the usefulness of the paper. The authors claim that sulfur is the major source of production problems in cement kilns. This is not the case. Sulfur is not a problem as long as the mole ratio of sulfur to alkali oxides is less than, or equal to, 1. This is attributed to the sulfur preferentially combining with the alkali oxides and not acting as a clogging agent due to the nature of the pre-heater vessels. Where the mole ratio exceeds 1, many methods are available to eliminate or reduce the clogging problems such as using the preheater bypass or a precalciner with up to 100% kiln gas bypass. In non-preheater kilns, the presence of sulfates is hardly ever a problem anyhow. Thus, the level of sulfur in the fuel and raw material that might cause problems is not a fixed amount, but is predictable. The purpose of the paper is to discuss the economics of conversion of kilns in the Indonesian and Phillippine cement industries from oil firing to coal firing. This economic analysis has very little to do with the sulfur levels. Thus, the section of the paper on sulfates is actually rather meaningless. The economic analysis is also rather meaningless because the authors have failed to follow accepted economic analysis procedures by including interest charges. They have fallen into the trap of confusing financing with economic evaluation. When the accepted methods of economic evaluation are used, the return on investment increases to about 35%. In conclusion, the paper is not a meaningful contribution, and sadly, quite erroneous.

Jan 11, 1982

By Thomas J. Ferree

A Reichert cone concentrator was installed in an aggregate plant circuit to recover the extremely fine "flour gold" known to exist in the deposit. The gold content of the bank-run gravels averages about 0.1 grams per ton (about $1.45 per ton), based upon $14.47/g (345O/ounce) gold. Tine Reichert cone, in conjunction with a conventional shaking table, re- covers in excess of 85% of the gold. A screen analysis of the gold particles shows 44.1% is -i4 microns (-200 mesh) in size. The plant operates with a two-man crew.

Jan 1, 1981

By Albert J. Shohet

The increasing scarcity of bentonite coupled with the potential of improved pellet strength have created a need for understanding the effects of thorough bentonite dispersion in taconite concentrate. It has been known for more than a decade that under certain conditions substantial savings of bentonite are realizable by using thorough mixing of bentonite as compared to poor mixing. Extensive work done by J. O. Zetterstrom l at the Univer¬sity of Minnesota indicated the effect of thorough bentonite mixing in a controlled laboratory testing. Commercial hardware was not available at the time Zetterstrom1 paper was published and not much work was done until R. L. Stone, J. M. Karpinsky2, did work with Muller mixing. Nabush 3 worked with the pugmill principle, and Hoogovens 4 and Robe River 5 used the fluidized bed principle. It was very hard to convince iron ore industries to improve the mixing of Bentonite and taconite for the following reasons: A. The initial capital investment for improved mixing equipment; B. The relatively low bentonite cost a few years back, which was not enough to justify new equip¬ment investment;

Jan 1, 1976

By R. D. Lopez

The concept of "heat recuperation" for the grate-kiln pelletizing system was first implemented at the Sydvaranger plant in Norway in 1974. Essentially, heat recuperation involves bringing hot air from the second stage of the pellet cooler to the drying zone to increase drying-air temperature to the maximum limit. 'Burner-fuel savings result from a lower kiln secondary-air mass flow at a higher temperature. Because of the fuel benefits achieved at Sydvaranger, heat recuperation has been included innumerous new grate-kiln plants built since then, and some older plants have been retrofitted for heat recuperation. In recent years there has been much confusion about the fuel benefits of heat recuperation. To answer some of the questions, a mathematical model was developed and plant tests were conducted at Minntac Phase 111. Minntac is a magnetite taconite concentrating and pelletizing plant, and Phase III has two indurating lines. This paper presents the results of the studies and plant tests.

Jan 1, 1980

The Bureau of Mines, U.S. Department of Interior, is currently investigating strata interactions associated with the mining of multiple coalbeds. Strata interactions due to subsidence are a result of undermining. This study involves underground investigations at a mine where room and pillar panels are actively being developed through an area of the Sewickley Coalbed which was subsided by undermining of the Pittsburgh Coalbed. Research to date indicates that undermining had little effect on upper mine pillar stability, but had a more severe effect on the development and maintenance of entries. Roof-to-floor measurements recorded over four times more convergence in entries developed over gob as compared with entries developed over support pillars in the lower mine. Consequently, this has led to major roof falls in the active panels, resulting in considerable production delays and supplemental roof support costs.

Jan 1, 1985

By Jacques LeBel

After a brief definition of asbestos and a review of regulations, the elements of asbestos dust control are outlined for various asbestos manufacturing industries. To prevent the asbestos dust to become airborne, the general procedure is to control the dust at the source by enclosure and exhaustion. The asbestos dust is collected and removed from the air with bag filter collectors of high efficiency (better than 99%) to ensure that the air emitted to the general environment or returned to the workplace are below the existing asbestos emissions regulations. Dust control systems should be regularly inspected and serviced to maintain at all times the efficiency and reliability of the equipment. Other effective procedures for asbestos dust control are also considered.

Jan 1, 1984

By F. C. Townsend

Centrifugal modeling was used to evaluate reclamation schemes of; (a) flocculants, (b) sand/clay mixes, (c) flocculants/sand addition and (d) capping with sand or sand/clay mix. Results suggest the hierarchy of effectiveness is: (d) capping, (b) s/c mixes, (c) flocculants/sand addition and (a) flocculants. Comparisions between currently used large strain consolidation computer programs reveal no difference between piecewise linear or finite strain calculations. Correlations between numerical and centrifugal model predictions reveal excellent agreement on magnitudes; however, rate predictions require improvement in laboratory input data and centrifugal scaling relationships.

Jan 1, 1987

By Rudecindo S. Silo

[The principal characteristics of the process is that when applied in medium or in great degree it only needs unelaborated raw materials such as sodium chloride and calcium carbonate, or of medium elaboration, as sulphur which is found in large quantity in the North of Chili; however, the may be also applied to small copper plants provided that low cost sulphuric acid is available as a substitute for sulphur, that is, in. those plats of small capacity where large additional installations are not needed for the regeneration of the leaching solution. The method itself comprises 6 stages, namely: crushing, leaching, washing; of solutions, precipitation, regeneration of the leaching solution and fusion.]

Jan 1, 1965

By R. K. Singhal, D. B. Stewart

A variety of equipment is being used for roadheading and tunneling in the mining industry. In United States drum-type continuous miners are used most extensively. In Europe and Canada mine roadways are driven mainly by roadheaders. The application of tunnel boring machines to coal mining is somewhat recent. Such machines are however being used on major projects driving main mine entries.

Jan 1, 1985

By D. D. Misra

Though computer simulation of coal beneficiation has been in use for nearly two decades the models used in the programs for the simulation of the crushing processes concentrate mainly on the size distribution. This paper suggests the use of an empirical equation approximating to the cumulative probability or error function for the simulation of the liberation process in coal crushing. The algorithm outlined here can be incorporated in a computer program already having a size distribution model.

Jan 1, 1989

By B. C. Haydon

Manganese silver ores are well known to be refractory towards cyanidation. Various processes have been suggested for improving the amenability of these ores. A sulfurous acid pre-leach is one such process. This paper traces the development and implementation of the sulfurous acid pre-leach process for the retreatment of manganese silver tailings at Tonopah, Nevada. An account of the test work program is presented, followed by a description of the operating plant and specimen metallurgical performance. The paper also summarizes the history of the Tonopah tailings, their mineralogy and makes reference to previous application of the sulfurous acid pre-leach process.

Jan 1, 1984