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By D. M. Thayer

Dry magnetic separation of feebly magnetic minerals and wet magnetic separators suitable for highly magnetic minerals have been commercially utilized for a number of years. The beneficiation of magnetic taconite ores by conventional, low intensity magnetic separators in combination with flotation is well established technology (1). However, over the past decade much effort has been expended in developing techniques for the upgrading of feebly magnetic oxidized iron ores (2,5). Such research and development have been stimulated by the limitations of the above-mentioned techniques in application to ores containing fine material; for example, ores requiring fine grinding for liberation or slime-forming ores such as the oxidized taconites or semi-taconites. Wet separation has allowed the reduction of the minim particle size that can be effectively treated from that allowable for high intensity dry magnetic separation. Consequently, the development of high intensity wet magnetic separation methods has received much attention in the search for an effective method to treat the weakly magnetic minerals. Recently Lawver and Hopstock (5) have provided an excellent review of the state of know- ledge of this technology. In addition, a valuable compendium of available laboratory and commercial devices is included. Nonetheless, the feasibility of the use of high intensity magnetic separation in a particular application is largly empirical in nature and requires experimentation to chazacterize the response of a particular ore. The purpose of this paper is to assess the incorporation of high intensity magnetic separation for the beneficiation of a non- sliming specular Michigan hematite. Particular attention is focused on the existence of a lower limit of particle size that can be treated effectively and is inherent in the use of these techniques. Additionally, a rationalization of various models presented in the literature is attempted.

Jan 1, 1979

By L. Alagha, M. Khodakarami

"This work focuses on the extraction of rare earth metals (REE) from aqueous solutions using stimuli-responsive polymers. Various parameters affecting the adsorption capacity of the polymer, including pH, polymer concentration and REE concentration were studied. The interfacial behavior of REE in aqueous phase before and after polymer addition and the mechanism of the formation of metal-polymer complexes and their properties were investigated using UV/VIS spectroscopy. Results obtained would help to test the feasibility of employing stimuli responsive polymers in pre-concentration of rare earth elements associated with other elements in mineral aqueous streams prior to solvent extraction. The pre-concentration step would reduce the amount of organic solvents used for extraction and striping compared to conventional extraction processes. INTRODUCTION Rare earth elements (REEs) have become the foundation for several technologies and have been used extensively in many industries such as modern electronics, green energy, health care and transportation and defense [1-4]. As these industry sectors are growing, the global consumption of REEs is going to increase dramatically. The current global REEs demand is estimated to be 105,000 tons of rare earth oxides per annum with an expected annual growth of 5.6% [5-6]. Rare earth elements are produced from different sources (primary and secondary). China is considered as the largest producer of REE in the world, providing about 90% of the world’s mine production while the United States is second with 4% of the mine production [7]. High global demand for rare earth materials results in shortage of such materials in the near future. It should be considered that the term ‘‘rare’’ earth is not an accurate term, since they are abundant in some extent in the Earth’s crust, but are typically finely dispersed. Thus, nowadays looking for more efficient and cheaper methods for extraction, separation and purification of low concentrated REE ores and exploring secondary sources are essential. Rare earth materials are found in raw condition and require treatment that involves several processes, which can consume money and time. Recent improvements in extraction and recycling technology have made the extraction of rare earth elements from rare earth sources more feasible in terms of cost and time [8-10]. Physical and pyrometallurgical methods require high-energy and thus high costs. Moreover, due to environmental problems, environmental protection systems are required, and they are not as adaptable as chemical processes [11, 12]."

Jan 1, 2016

By R. H. Parker, B. A. Wills, D. G. Binns

Introduction With the increasing need to mine lower grade ores, high energy-related costs in comminution are of major concern. Very fine grinding is needed to liberate the fine mineral particles in these ores. Not only is this expensive, but it leads to greater losses in slimes. Gravity concentration techniques become unacceptably inefficient for particle treatment below 25 to 50 µm, and even flotation fails in the ultra-fine range. Ore from the South Crofty mine near Redruth, Cornwall, was subjected to thermal pretreatment in the hope that differential thermal expansion of the minerals would lead to intergranular cracking, and thus enhanced liberation. South Crofty ore contains about 1% tin as cassiterite associated with a complex assemblage of quartz, chlorite, tourmaline, and hema¬tite in granite and slate. Recovery of cassiterite can be as low as 70%, as overgrinding occurs during stage reduction of the material to 1 mm (primary grinding) and 180 µm (regrinding). Cassiterite is present in grain sizes ranging from submicroscopic to + 10 mm, the bulk being predominantly in the range of 50 µm to 3 mm. Experimental Flat polished sections of the ore were photographed in reflected light using a Vickers M17 microscope. The sections were then heated in a Carbolite LMF4 muffle furnace. The heating rate and temperature were monitored by a thermocouple immersed in the bed of particles. Previous work by Sherring (1981) on the same material showed that a 55% reduction in grinding resistance occurred when the material was rapidly heated and cooled through the quartz inversion tem¬perature (573°C), where a volumetric expansion of 0.86% occurs. The samples were therefore heated to 650°C at 26°/min, and were then water-quenched to room temperature. The effect of heat treatment on the mineralogy and fracture network was assessed by examining the same area after treatment. Results Most of the sections examined showed that extensive transgranular cracking occurred as a result of heat treatment. Such cracking, although weakening the rock and hence reducing the work index, would in no way enhance the liberation of the cassiterite from the host minerals. Intergranular cracking, which would lead to enhanced liberation, was difficult to discern under reflected light, but there was evidence of such cracking in some of the sections examined. Figures 1 and 2 are examples of typical sections before and after heat treatment. Figures la and lb - show cassiterite in hematite, Fig. lb, in normal incident illumination, illustrating clearly that after heat treatment, transgranular fracture is evident in the cassiterite. If any intergranular cracking has occurred, it is not evident in this photograph. Figures 2a and 2b show cassiterite in quartz before and after treatment. Fracturing in both minerals after treatment can clearly be seen, and there is evidence of intergranular fracturing. However, protruding "arms" of cassiterite are severed by transgranular fracture. Subgrains of quartz have also been isolated by trans¬granular fracture. There is major transgranular cracking across the wider sections of cassiterite. Discussion The effect of rapidly heating South Crofty tin ore to 650°C, followed by water-quenching to room temperature, has been studied by observing the fracture networks in mineral grains. The choice of 650°C as the heating temperature was influenced by the work of Sherring (1981), who found considerable reduction in grinding resistance after thermal pretreatment of the same ore. Scheding et al. (1981), however, have shown, by means of a crude calculation, that heat treatment cannot be justified solely on the basis of reduced grinding costs. The cost of heat treatment far out-weighs that of grinding, resulting in the combined costs for heat treatment being over six times that for grinding unheated material. The most economically attractive aspect of heat treatment is the possibility of enhanced liberation of the valuable mineral due to increased intergranular rather than transgranular fracture. Grinding costs would be greatly reduced by improved liberation at coarser sizes, and the costs of ancillary processes, such as dewatering and tailings disposal, would be reduced. The most significant economic effect, however, would be in improved metallurgical efficiency. Improved liberation would increase concentrate grades, and recoveries would be higher, particularly in the case of ores where high slime losses are produced due to the excessively fine grinding required to produce adequate liberation. Manser (1983) has shown that only a 1% increase in tin recovery, at the same concentrate grade, would be sufficient to offset heat treatment costs on South Crofty ore. Heavy liquid analysis and shaking table separation have been used to evaluate the effect of thermal treatment on processing (Binns, 1984; Scheding, 1981; Sherring, 1981). However, treatment of unheated and heated samples of South Crofty ore, ground to the same product size, by such methods have shown little evidence of improved metallurgical efficiency after thermal pretreatment. The large amount of transgranular fracturing revealed by reflected light studies helps to explain this lack of improvement. Nevertheless, from the evidence of some of the fracture work, it is surprising that no improvements in metallurgical

Jan 1, 1988

By R. L. Grayson, R. C. Althouse, L. A. Layne, M. J. Klishis

Introduction The researchers combine a detailed microanalysis of accident data and roof bolting field observations to develop new and meaningful risk indices for the roof bolter job classification. These indices incorporate information that has not been used in traditional measures of risk. The significant parameters that these new indices incorporate are the total number of accidents occurring within a particular work activity, the amount of time a miner usually spends performing the work activities in which the accidents occurred, and the average severity of the injuries incurred in the accidents. Being specific for a particular job, they aid comparisons among work tasks performed in the bolting cycle. The indices' specificity provide anew basis for reexamination of training and operational development of miners for specific jobs. Microanalysis The roof bolting microanalysis was performed by examining the empirical data and written summaries for all bolting accidents reported in the West Virginia Safety Information System (WVSIS) for the years 1983-1987. The WVSIS database contained more than 15,000 accident narratives. The microanalysis of bolting accidents covered 2083 reported personal injuries (RPIs). This large number of injuries, performed at an industry-wide level, provides a broad picture of the nature of roof bolting injuries. The analysis focuses on the activities associated with the cyclical routines of the roof bolting process. For this analysis, roof bolting tasks were categorized into work routines based on the sequence of work activities performed by roof bolters. Four work routines were identified within the overall bolting cycle. And these routines comprised 70% (1217) of all roof bolting accidents at the face. The routines include: face area preparation; tramming, positioning and setting the automated temporary roof support (ATRS); drilling holes; and installing bolts. The types of injuries most common within a routine and the average severity of those RPIs were also identified. Table 1 gives a breakdown of the percentage of accidents by the bolting routines. Another 19% of the roof bolting RPIs cannot be classified in a particular routine because of the vagueness of the narratives. However, it was clear that these accidents involved drilling holes or installing bolts. These results are consistent with previous research. It identified drilling and bolting, followed by tramming, as the most frequent sources of injuries (Helander, Krohn & Curtain, 1983). [Table 1-Accidents by bolting routines at the face Accidents Bolting routine Percent Number Face area preparation 18.2 222 Tram, position, ATRS 23.3 284 Drilling holes 34.2 416 Installing bolts 24.2 295 Total 1217] Roof bolter operators were involved in 82% (1706) of the 2083 accidents. The RPIs incurred by classified roof bolters occurred largely on working sections (95%) and at the face area on the working section (75%). Approximately 23% of the 1706 roof bolter RPIs occurred during tasks which were not part of the bolting cycle. These included belt utility work, providing supplies and section moves. Miners in other job classifications who were involved in roof bolting accidents accounted for the remaining 18% of the 2083 roof bolting RPIs. Among this group were section utility men and general laborers (41%), section machine operators (38%), and electricians and mechanics (11%). Another 9% were performing on-section supervisory functions. Face area preparation Face area preparation work led to a substantial number of injuries, although it accounted for the lowest number of injuries of the four routines. Several tasks were associated with the injuries incurred in this routine. Table 2 shows the percentage of injuries associated with face area preparation. [Table 2- Accidents occurring during face area preparation 26.6% Setting temporary and supplemental supports 24.8% Scaling roof and ribs 21.2% Installing ventilation devices 15.8% Handling supplies and materials 11.6% Miscellaneous tasks] The injuries commonly involved overexertion and 'struck by' falling coal and rock. Handling materials often led to overexertion injuries. Nearly 66% of the scaling-related injuries and about 33% of the injuries incurred during the setting of temporary supports involved 'strikes' by equipment, materials, top and rib. Tramming, positioning and setting the ATRS During tramming operations, workers were often struck by components of the machine and hit by loose roof and rib. Poor footing also led to slips, sometimes against the moving machinery. Table 3 gives a breakdown of accidents in this routine. Control levers were a significant problem in this routine.

Jan 1, 1993

By R. S. Datta

Regulations on sulfur emissions have caused the electric utilities to evaluate various methods of meeting them. Low-sulfur, low-volatile coals can be burned alone or blended with high-volatile? coals. This option may be a cost-effective alternative to the installation of scrubbers on coal-fired combustion systems. In some blending applications, problems may be encountered because the low sulfur content and high calorific value of the low-volatile or other clean coal allow it to economically upgrade extremely poor quality coal. While this is done to meet specification purposes, actual boiler performance may be dominated by the poorer quality coal. Thus, the reputation of the low-volatile or other clean coal suffers. As a result of laboratory and full scale testing, conducted over the last five years, a model has been developed to evaluate the relative combustion characteristics of various coals and coal blends. A database for different coals has been gathered. This paper presents the model and shows how power plants and coal mine personnel can use it to evaluate their coals from different sources. It can be used to blend two or more coals from different sources so that thee blended coal has flame stability characteristics which are similar to those of the baseline coal.

Jan 1, 1991

By Lennart Bergstedt

Boliden Aktiebolag, the prime base metal mining company of Sweden, annually treats some 11-12 million tons of sulphide ore in nine concentrators. Its first flotation plant was established in the early thirties for the complex Boliden ore. Since then, the number of mines and plants has gradually increased. Considerable effort has been devoted to the development of all phases of the technology of ore prospecting, mining, milling, smelting and refining. This has been necessary due to the complex and low-grade ores. The development of concentration has mainly been directed toward the arts of grinding and flotation. During the forties and fifties, open-stage fine crushing followed by two-stage grinding in rod and ball mills with cyclone classification was successfully introduced and used throughout the plants. The increasing cost of grinding media led to studies of an autogenous grinding technique as early as 1945. Basic test work commenced during the early fifties, with a. number of ores tested for pebble grinding. In 1956 a new pilot plant was built, with equipment installed for the continuous testing of pebble grinding and, later, for primary autogenous grinding of run-of-mine ore. The intensive test work showed that pebble grinding was feasible in almost all of the plants. At this early stage of testing, however, the merits of autogenous grinding could not be fully evaluated. As a. consequence the rod mill/ball mill grinding systems were converted to rod mill/pebble mill systems whereever possible. In parallel with the converting of grinding systems the development work concerning the merits of autogenous grinding went on. The major result

Jan 1, 1977

By Someswar Kesh, Gavin J. Faulkner

Introduction Powered supports in longwall mines essentially operate under two conditions: 1. Static (non-working) 2. Dynamic (working) There is a variety of static conditions which may exist in a mine. These may include a temporary stoppage of work for a few minutes or stoppage of work for days because of strikes or lack of demand for coal. Studies on pressure behavior were made during brief stoppages of work (few hours) as well as a long stoppage of work (two weeks). In dynamic conditions, the pressure variation on supports is expected to follow a certain pattern. However, in many cases that does not happen. Pressure-time characteristics of each leg of the supports were studied. This paper discusses the research findings of the studies made on powered support pressure behavior under both static and dynamic conditions. Studies on Pressure Variation in Static Conditions It has been noted (Brenner et a1.,1975) that the pressure distribution on mechanized longwall faces is random in nature and changes with time and position along the face. Figure 1 and Figure 2 show the distribution of support pressures for the front left leg of supports in two different mines. In both the cases it can be seen that the variation of the support pressures along the face is indeed absolutely random. However, the nature of randomness in both the cases is very different. The data in Figure 1 (Mine 1) is much more random than the data in Figure 2 (Mine 2). This can also be seen from the standard deviation of the data (81 26 Pa versus 4000 pa) . By comparing the results of ten other sets of data the same conclusions can be arrived at. Since it was found that in Mine 1, the pressure on support legs was extremely random, batches of ten supports each were taken and the mean pressure plotted. This mean pressure line is shown in Figure 1. It was expected that this would smooth out the variation due to individual props and reveal a better picture of the support pressure distribution. For the case cited, the standard deviation is lowered to 5043 Pa from 8126 Pa which indicates that, on an average, the supports tend to take loads more uniformly, however for individual support legs the load on the supports is very random. The significance of this is that, if the supports are selected on the basis of theoretical load calculations alone (based on the roof block concept or some other concept), only the mean value of the support load required will be known. In the case of a pressure distribution with a high standard deviation, some supports selected on the basis of mean value will fail, while others will be supporting low loads. Hence, for proper selection of support capacity, it is not only necessary to know the average load but also the expected nature of the distribution. The only reason why the supports are supporting the roof in Mine 1 well is the fact that the yield load is above the highest level of load attained. If Mine 2, which has a much higher average load, had the same sort of standard deviation, the loads on the individual legs could easily exceed the support load. For example, if the pressure distribution followed a normal curve and the mean pressure was 36,700 Pa with a standard deviation of 9000 Pa, it is possible that 67% of the supports will be within the pressure range of 36,700 plus or minus 9,000 Pa, or 22,700 and 45,700 Pa. Many of the supports could then exceed the yield load of 40,000 Pa. It is clear then that the mean load is an inadequate criterion for estimating support resistance. The nature of the problem is statistical and such an analysis should therefore be made as far as pressure data are concerned. Statistical Analysis of Pressure Data. In general, in longwall mining and other geo-mechanics studies, it is extremely difficult to specify the exact nature of the distribution of a variable. Under these circumstances, probability

Jan 1, 1986

By C. J. Im

As market conditions become more competitive, U.S. coal producers are faced with the increasing need to find methods of lowering production costs in their operations. One area worthy of investigation is the additional recovery of fine coal that is normally discarded in the preparation plant waste stream. United Coal Company is currently assessing the capability of the Deister Flotaire, a new application of a column flotation cell, to recover additional, saleable coal from preparation plant waste. This paper presents the technical data obtained to date. Work with a 0.76 m (30 inch) diameter cell installed in a plant shows considerable coal recovery is possible from a nominal 0.595 run (28 mesh) X 0 refuse stream.

Jan 1, 1986

By John W. Leonard

The typical tunnel boring machine in use on today's tunnel project uses 15'2-inch diameter cutters at a nominal load of 45 kips per cutter. The cutterhead rotates at a speed of 120 over D revolutions per minute where D is the cutterhead diameter in feet. The typical machine operates about 50 percent of the time. Delivery of a new machine takes 12 to 15 months. The TBM is a specialized tool and not applicable to all jobs. There is a range of job conditions that limit the economic use of a TBM. The machine should provide a thrust per cutter at least twice the compressive strength of rock in psi and up to three times the rock strength is required if in "granitic rocks." The rock strength in psi must be at least three times the cover above the tunnel in feet or the machine may stick. For example, 2,500 feet of cover requires at least 7,500 psi rock to avoid problems. When the RQD's over a significant distance of the project are less than 50, the TBM becomes support-bound. When a large percentage of the time is spent waiting while the ground is supported, the TBM cannot perform well (See Figure 1). This problem increased inversely with tunnel diameter. We are finding that increasing the RPM's does not significantly increase progress. The increased RPM results in a substantial decrease in cutter penetration. The explanation for this decrease appears to be vibration, cutterhead drag, or lack of stiffness in the cutterhead (See Figure 2).

Jan 1, 1981

By Ingemar Marklund

INTRODUCTION Today at LKAB, the primary mining method used is sublevel caving, independent of ore-body size, shape, dip, and geographical location. The mining of small ore bodies with relatively flat dip involves a lot of develop¬ment work, a large initial investment, big dilution, and large ore losses. In 1975 serious investigations into other possible methods, especially for those small ore bodies which do not lie in direct contact with other ore bodies, began. Different methods were studied and a number of in¬teresting ideas came from Russian mines. An analysis was made of one of the proposed methods, sublevel benching, in which mining is carried out from raises, and which uses the throw from the blast to place the broken ore. However, there were reservations about working in big open rooms at great depth, and doubts about the possibility of getting acceptable drilling results in an acceptable environment from raised platforms. Testing was therefore delayed for a time. However the investigations continued and in the fall of 1976, it was decided to try vein mining in a small ore body (Indian) (see Figs. 1 and 2). MINING RESULTS The results are now available from the mining of Indian and are as follows: Ore plan area 1200 m2 Level height 50 m Ore volume 60 000 m3 Ore tonnage 270 000 t Iron content 62% Dip 0.69 rad (40°) Ore thickness 14 m Ore compressive strength 26 MPa Ore C-factor (fracture density) 0.21 Hanging wall compressive strength 28 MPa Hanging wall C-factor 0.27 About 40 000 m of the holes were drilled by ring drills with an average productivity of 85 m/shift; loading of a round (2 rows, 750 m of hole) was usually accomplished in one shift. The total ore recovery was about 270 000 t with an average iron content of 59%, and three to four shifts were required per round (4700 t) to load. The principal of "throwing" the rock during blast¬ing worked well, and no ore remains on the footwall. ROCK CONTROL Before mining commenced at Indian, it was dis¬covered that the rock strength in its vicinity was rela¬tively poor. Since the ore is flatly dipping, and mining was to take place with open rooms, the strength of the hanging wall was of great importance. A supplementary program of exploration drilling was done during the fall of 1976 in order to assess fracture spacing and rock strength from the drill cores. Results showed that the strength of the hanging wall was low because of the many fractures. A decision was made to try cable bolting of the hanging wall by extending some of the ordinary blast¬holes. Cables (20 mm diam) were grouted in these larger holes and the rest of the hole was then utilized for ordinary loading (see Fig. 2). About 100 bolts altogether were installed and, to date, only small falls have occurred and the hanging wall is holding. Its exposed area is 2100 m2 DEVELOPMENT WORK On the 350-m level, four crosscuts (two per mining room), one longitudinal drift, and two elongated drifts (one per raise) for the raise climber were developed as shown in Fig. 1. Of the 230 m of drift, 60 m were in ore and the rest in waste. For mining room A alone, a total of 115 m of drift were needed with 30 m in ore. The drift dimensions were about 4.5 X 5.5 m. In the drifting on the 300-m level, two connecting drifts were driven from one existing transport drift forward to the respective raises. The total development was 93 m, all in waste (Fig. 1). The two raises having a length of 80 m each were driven at an angle of 0.68 rad (39°) from the hori¬zontal. They were driven along the footwall contact with a separation of 30 m and had an area of 10 m2. LONG-HOLE DRILLING The drilling of long holes with the ring drilling unit began in the fall of 1977 in raise A. The equipment consisted of an Alimak raise climber, equipped with a lower platform from which the charging was done.

Jan 1, 1982

By NARAYAN K. BHAGAT, ARVIND K. MISHRA, Aditya Rana, M. M. Singh

The paper presents an innovative directional controlled blasting (DCB) technique for excavating unstable slopes along the Konkan railway in India without disturbing the traffic. New empirical formulae for small geometry blasting to estimate burden and throw were 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 was changed from 80–82 to 45–47 degrees, and possible wedge, planar and toppling failures were eliminated. No incident of rockfall and slope failure along these slopes has been observed after the stabilization.

By William A. Rehrig

During recent years, there have been frequent investigations on the geochemistry of biotite. This work has emphasized, and no doubt was prompted by, factors related to mineral exploration, particularly for those deposits of the porphyry copper type. A large number of papers have been devoted to the obtainment of trace-element data from plutonic biotite. These data have been used to extract primary geochemical information pertaining to the parent magna and the hydro-thermal system which together give rise to porphyry copper deposits. In two early works, Putman and Burnham (1963) and Parry and Nackowski (1963) analyzed base metal contents of ferromagnesium minerals, primarily biotite, from various barrenfmineralized plutons of the Basin and Range Province. One similar conclusion from the two studies was that the Cu, Zn, and Pb contents of magmatic biotite served as indicators of ore development within the granitic rocks genetically related to ore. This theme was then pursued vigorously by a number of authors until the present time. Results, however, have been contradictory, as reviewed by Coope (1974). For example, studies by A1 Hashimi and Brownlow (1970), Lovering et a1 (1970), Putman (1972), and Graybeal (1973) continued to support positive primary relationships between the trace-metal content (usually copper) of biotite and associated sulfide mineralization. By studying biotite in close spatial association will sulfide occurrences, Putman (1972) and Graybeal (1973) developed models for base metal partitioning in magmatic systems based on their discovery of inversely distributed copper between silicate (biotite) and sulfide phases. Conversely, research by Bradshaw (1967) and Blaxland (1971) plus the recent findings of Parry and Jacobs (1975) and Kesler et a1 (1975) found little if any evidence to support a correlation of biotite base-metal or

Jan 1, 1976

By S. J. Schatzel, J. D. Addis, V. Gangrade, L. L. Chasko, C. A. Hollerich

"A ventilation study using tracer gas was conducted at a western US coal mine. The objective of the study was to evaluate the movement of longwall face air exchanges between the face and worked-out area and to document the presence or absence of face airflow pathways between these locations. The mine operator uses a bleederless longwall ventilation system with a back return and a blowing mine ventilation system. The study was conducted on an active panel and included both underground and surface monitoring sites. The study used sulfur hexafluoride (SF6) released as a slug on the longwall face and in the front of the gob inby the face. The velocity of the tracer gas movement in the gob was 0.019 m/s (3.7 fpm). The rate of movement for the overall tracer gas slug averaged about 0.0091 m/s (1.8 fpm). A separate tracer gas test initiated with the release of SF6 into the legs of the first shield showed the existence of more than one pathway of face air in the general direction from the headgate towards the tailgate corner. Maintaining adequate ventilation air on longwall faces is important for worker safety and for the dilution of methane emitted from the face and caved gob. A more detailed characterization of longwall system air and gas movement allows a mine to better assess its ventilation design for controlling gas on the face and in the gob. INTRODUCTION During a series of stakeholder meetings, a near unanimous concern was expressed related to maintenance of effective ventilation airflow on longwall faces. There was a strong interest expressed regarding the impacts of face length and caving characteristics on the ability to maintain adequate face airflow levels. This study seeks to understand and improve a mine operator’s ability to monitor and control ventilation and methane concentrations across the longwall face by assessing airflow variations in terms of roof caving characteristics and face length. National Institute for Occupational Safety and Health (NIOSH) researchers seek to identify flow paths of ventilation airflow on longwall panels with variations in roof caving characteristics and longwall face lengths. The project research aims to develop an understanding of the role of gas dynamics in variable broken rock, interactions with ventilation airflow and methane concentrations across the longwall face area. The research also addresses continuous ventilation monitoring along the shield line to provide early detection of gas exchanges between face and gob, and thus improve a mine operator’s ability to monitor and control ventilation and methane concentrations around the face. In the context of this paper, the longwall gob refers to the pillared area behind the shields. This work reports on results from one project task to conduct field based research using a tracer gas to describe face air and gob area gas movement in a bleederless longwall ventilation system that incorporated a back return."

Jan 1, 2017

By A. K. Sheik

A small underground mine in northern Idaho was used as an experimental facility for studying the role of water in the analysis of mine opening stability. Laboratory test data, borehole measurements and geological mapping assisted in constructing a coupled finite element model of the rock mass. The model was used to make quantitative estimates of: 1) rock stress, strain and displacement changes, and 2) water pressure and flow velocity changes that occurred during historical mining. Stope cuts and wet fills were included in the analysis. Stability of crown pillars between stopes was of particular interest. The relatively low effective stress and high rock mass strength showed that water pressure changes were of no great consequence in agreement with experience. The computational tools used in the analysis may be applied to other wet mines with appropriate changes to site-specific input data.

Jan 1, 1999

By M. Karmis

The determination of the surface area which may be impacted by underground mining is an important consideration in subsidence engineering. A number of parameters have been utilized to identify this influence area, most notably the angles of draw and break. This paper discusses the implications of selecting an appropriate "damage angle" for defining a potential damage area due to mining, within which surface structures may be impacted Furthermore, the influence of both mining and geological factors, such as mining width and depth and overburden lithology, on this angle is examined, using the Surface Deformation Prediction System (S.D.P.S) software. Finally, recommendations are given for applying these concepts in subsidence control.

Jan 1, 1994

By Barry Hunsaker

I cannot help but be reminded, as I stand here today, of my boyhood years and heritage, both of which are inextricably tied to this very valley and are, in a: bay, apropos of the theme of this Keynote Session. I remember the scarcity of people. For example, 48, 000 in Phoenix and 436, 000 in the State in 1930 and 65,000 in Phoenix and 499, 000 in the State in 1940. Today Phoenix boasts a population of over 500, 000 and the State a population of 1, 500,000. I remember the' lack of facilities, goods and services, including gas and electric power. I recall stories of my grandfather who had the freighting or teaming contract to haul all the supplies and generating equipment, by wagon mind you, up the old Apache Trail to Roosevelt Dam. . .a three day trip from the nearest railhead which, by the way, was not Phoenix, for it did not merit a railhead on those days, but the village of Maricopa approximately 25 miles south of Phoenix. I remember chopping firewood, heating wash water on an open fire in a No. 3 wash tub at the rear of the house and doing my school homework by the light of a kerosene lamp just 25 to 35 years ago within what is now a large metropolitan area

Jan 1, 1965

By T. E. Mccandless

Detrital minerals of mantle origin have been found in the Green River Basin of Southwestern Wyoming (Figure 1). Pyrope, pyrope-almandine, and chrome-bearing clinopyroxene and orthopyroxene occur on the antmounds of Pogonomyrmex Occidentalis. These antmounds are a common sight throughout the western U.S., and were used by early paleontologists as fossil mammal collection sites. The mound is composed of two types of material; (1) soil excavated during the digging of the nest below the surface (maximum depth 3 meters) and (2) larger grains collected from the surface and placed on the mound, probably to prevent erosion. The selection of grains to armor the mound seems to be based largely on the mean diameter of the grain, with other factors such as weight, sphericity, roundness and composition in decreasing order of importance (Wheeler and Wheeler, 1963). The minerals of this study happen to fall within that mean diameter range. The garnets are found on the mounds as 3-6 mm diameter subrounded to subangular grains. The pyroxenes are also 3-6 mm in diameter and are most commonly subangular laths. The minerals themselves show evidence of abrasion, such as frosted surfaces and rounded edges. Orange peel texture is present on some of the grains, which suggests that kelyphitic rims were once present, but have been removed. Also, no similar minerals were found in sieved soil auger samples to depths of 2 to 6 meters taken near the antmounds. These observations imply a distal rather than proximal source for the minerals. Lamproitic rocks of upper mantle origin are found bordering the "Green River Basin. The best known are the Leucite Hills, a series of ultrapotassic volcanics found on the north flank of the Rock Springs Uplift (Carmichael, 1967). Minor occurrences of mica peridotite and ultrapotassic volcanics are also found on the western end of the Uinta Mountains in Utah (Best, et al, 1968). None of these rocks are known to contain minerals like those on the antmounds. The garnets are comprised of purple chrome pyrope, red chrome pyrope-almandine, and pink to orange low chrome pyrope/pyrope-almandine. Oxide statistics for these groups are summarized in Table 1. These three groups are distinct when their Cr203 values are plotted against magnesium number (Figure 2). The chrome pyropes are highest in Cr203 (1.28-2.66%) and have high, restricted Mg number values (0.781-0.826). The chrome pyrope almandines have similarly high Cr203 (1.29-2.62%) but a lower Mg number range. The low chrome pyrope/pyrope-almandines have the lowest Cr203 content (0.09-0.73) and the widest Mg number range (0.534-0.836). Using the classification scheme of Dawson and Stephens (1975) the chrome pyropes are most similar to garnets derived from kimberlite and from garnet lherzolite, websterite, and harzburgite xenoliths in kimberlite. The chrome pyrope-almandines and the low chrome pyrope/ pyrope almandines are similar to garnets derived from eclogite nodules (Figure 3).

Jan 1, 1984

The paper describes a practical methodology based on numerical modelling for the prediction of progressive failure of underground mine stopes with cable bolt support. A case study of the 40-761 stope at Kidd Creek Mine, Timmins, Ontario is presented. The stope was mined by blasthole open stoping in 1986 as a lead stope within a mining block between 36-1 and 38-2 levels. Initially designed to peak at 38-2 level, the stope back caved upwards approximately 40m towards 36-1 level. Cable bolts were installed above the stope from 36-1 level after caving had begun. The stope eventually stopped to cave when the falling rock mass encountered a sufficiently high cable density. The paper provides details of the geomechanical data and modelling results of the case study. Numerical modelling is carried out using a finite element package known as e-z tools which has, among other things, a special cable/rockbolt simulation capability.

Jan 1, 1994

By Ronald M. Hays

During the twentieth century the Mesabi Range in northeastern Minnesota has been the major source of iron ore for American steel mills. The Mesabi Range extends along an iron bearing outcrop known as the Biwabik formation. This outcrop is some 120 miles long and up to three miles wide. The formation is from 200 to 650 feet thick and dips 5 to 10 degrees to the southeast. The iron content of the original Biwabik formation was quite uniform, but much of this formation has been subjected to alteration. Thus major variations in iron content and mineral composition exist both laterally and vertically in the present Biwabik formation. The remaining unaltered iron formation includes the magnetic taconites which are commercially processed by Reserve Mining, Erie Mining, Eveleth Taconite, National Steel, Butler Taconite, and U.S. Steel. These six operations have a capacity of 32 million tons. per year of iron pellets and have produced approximately 170 million tons of pellets since 1956. Capacity for another 30 million tons per year of pellets is being considered by the industry for the next ten years. This increase will be obtained by construction of new plants and expansion of present operations. Thus by 1977, it is estimated that approximately 60 million tons per year of iron pellets will be produced from magnetic taconite on the Mesabi Range. The portions of the Biwabik formation subjected to intense alteration resulted in the direct shipping, wash and heavy media ores. Since 1892, 1.8 billion tons of direct shipping ore have been shipped from the Mesabi Range. This ore has been the major source of iron for American steel mills. In 1966, 12 million tons of direct shipping ore and 18 million tons of wash, jig, heavy media and other gravity concentrates were shipped from the Mesabi Range. These products generally have poorer physical and chemical properties than the iron pellets produced from magnetic taconite. It is predicted that the shipment of these materials will decrease from about 30 million tons in 1966 to 6 to 8 million tons in 1975. This decrease will be due to depletion of reserves and poor quality product.

Jan 1, 1968

By B. H. Kim

The use of a 25.4 mm diameter water-only hydrocyclone for separating fine-size (-150 µm), high-density (>4.5 g/cc) particles from similar-size, lower-density (2.65 glee) particles was investigated. It was found that a single-stage separation recovered nearly 90% of the heavy minerals (7.0 g/cc), while achieving a volume reduction of about 70%. The use of dense medium in a single stage improved volume reductions to 90% at similar heavy mineral recoveries. In addition, computer simulations of a three-stage circuit were performed, which showed that recoveries approaching 97% with volume reductions of 95% were possible.

Jan 1, 1998