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By Igor V. Nazarov

INTRODUCTION Tens of deposits in' coal basins of Western and Eastern Siberia are being open-mined with the technology of overburden rehandling into the mined out space of pits by excavators - draglines. This technology is preferred due to the high productive Russian equipment, comparatively low cost, and relatively small Influence exerted on the environment. The basic tendency of the last decades is the aspiration for enlarging the sphere of the' technology application for development of gentle and steeply dipping seams with the use of blast destruction of overburden and with combined winning of loose alluviums and rock partings. Complex technological schemes with multiple rehandling are widely used. In addition to it, stripping workings, intermediate storage of mineral resources, placing of transport dumps, and engineering construction can be realized with the use of draglines. More than ten of dragline faces are simultaneously exploited in some pits, and the rehandling coefficient ranks from 0.1 to 8. Here, conditions of winning and placing (position of a face and existing dump) are varying in space and time; this leads to the necessity of reconsidering the order for conduction of works (schemes) within the current designing. By now, in order to this, it is required to realize complex engineering calculations by graph-analytical methods and definite experience. The need is available for automatization of this process by creation of the corresponding software. The software is to: • Orient on the active pits and use information on the existing geology, equipment, and situation in mine. • Calculate the existing schemes more operative and on a higher qualitative level to model their development in a pit space with parameter optimization. • Predict situations demanding to change the existing technology for mining operations and recommend optimal technological schemes. METHODOLOGICAL FOUNDATIONS FOR MODELING The problem is not new, it has more than fifty year-old history of creating mathematical models algorithms, and programs in Russia, Ukraine, USA, Canada. Australia and other countries with the developed mining branch. Conventionally, all population of programs can be divided into three groups depending on goal and interval of prediction: • Long-term programs used in designing and intended for substantiation of the nine capacity, selection of mining system, establishment of limits for applying the technology of overburden rehandling. Mid-term programs used in current planning for calculation of technological schemes and certificates of face in active mines.INTRODUCTION Tens of deposits in' coal basins of Western and Eastern Siberia are being open-mined with the technology of overburden rehandling into the mined out space of pits by excavators - draglines. This technology is preferred due to the high productive Russian equipment, comparatively low cost, and relatively small Influence exerted on the environment. The basic tendency of the last decades is the aspiration for enlarging the sphere of the' technology application for development of gentle and steeply dipping seams with the use of blast destruction of overburden and with combined winning of loose alluviums and rock partings. Complex technological schemes with multiple rehandling are widely used. In addition to it, stripping workings, intermediate storage of mineral resources, placing of transport dumps, and engineering construction can be realized with the use of draglines. More than ten of dragline faces are simultaneously exploited in some pits, and the rehandling coefficient ranks from 0.1 to 8. Here, conditions of winning and placing (position of a face and existing dump) are varying in space and time; this leads to the necessity of reconsidering the order for conduction of works (schemes) within the current designing. By now, in order to this, it is required to realize complex engineering calculations by graph-analytical methods and definite experience. The need is available for automatization of this process by creation of the corresponding software. The software is to: • Orient on the active pits and use information on the existing geology, equipment, and situation in mine. • Calculate the existing schemes more operative and on a higher qualitative level to model their development in a pit space with parameter optimization. • Predict situations demanding to change the existing technology for mining operations and recommend optimal technological schemes. METHODOLOGICAL FOUNDATIONS FOR MODELING The problem is not new, it has more than fifty year-old history of creating mathematical models algorithms, and programs in Russia, Ukraine, USA, Canada. Australia and other countries with the developed mining branch. Conventionally, all population of programs can be divided into three groups depending on goal and interval of prediction: • Long-term programs used in designing and intended for substantiation of the nine capacity, selection of mining system, establishment of limits for applying the technology of overburden rehandling. • Mid-term programs used in current planning for calculation of technological schemes and certificates of face in active mines.

Jan 1, 2002

By Pascale St-Germain, Michael Li, Lionel J. J. Catalan

The oxygen flux into water-covered reactive sulphidic tailings is required for evaluating the long-term performance and predicting the quality of engineered shallow water covers. This information can also serve as an input to the design of water covers for reactive tailings decommissioning. This paper describes some dissolved oxygen (DO) profiles recently measured in the field across the tailings/water interface and oxygen fluxes derived from these profiles. These fluxes are compared with previously reported flux values from similar field measurements. Techniques and pitfalls for measuring DO profiles and calculating oxygen fluxes are also discussed.

Jan 1, 2000

By Y. C. Kim

M.S. Azun The article unfortunately has some serious flaws negating the usefulness of the paper. The authors claim that kriging (linear kriging) is superior to other methods such as polygon. This is not the case. The deterministic methods (referring to polygon, ID, IDS, etc.) should not be compared with the statistical methods (linear kriging, disjunctive kriging, etc.) because they differ not only in their assumptions, but also their tools used in estimation. Thus, they give different results. By comparing the results of deterministic methods and linear kriging, the authors conclude that polygon, IDS, and linear kriging give an average thickness close to the value obtained from realizations. Besides, they argue that linear kriging gives less variance than the others (Fig. 2). The first conclusion is expected, due to the well-known central limit theorem. The second is the result of the statistical properties of different estimators. Thus, the comparison presented in the paper is rather meaningless because the authors have failed to follow basic statistical procedures. The numerical results of different approaches to an estimation problem cannot be generalized. In conclusion, the paper is not a meaningful contribution, and sadly, quite erroneous. Reply by E.Y. Baafi and Y.C. Kim The two key contentions made by M.S. Azun regarding the paper are: deterministic methods should not be compared with statistical methods and, hence, the results are erroneous; and the paper does not make a meaningful contribution. The first contention is an academic one. Most practitioners in ore reserve estimation could care less whether the tool they are using falls under a deterministic or a probabilistic method, so long as it gives the desired answers. It is probably for this reason that, whenever any new tool such as linear kriging is introduced to practioners in the mining industry, they would like to see a comparison test performed between the existing and the more familiar tools with a new and "strange" tool. Only when a new tool passes such a test, the generally conservative mining community would begin to utilize the new tool to its advantage. The second contention, is a subjective one. The final judgment should be left up to the readers within the mining community. If kriging is accepted more among practitioners, partly due to the paper, then the paper makes a practical contribution to the state-of-the art in ore reserve estimation. Finally, M.S. Azun entirely missed two main points in the paper. The first is that any comparative study of kriging vs. Polygonal or IDS should be performed for local estimation and not for global estimation. In the past, kriging was frequently compared with other existing tools in its performance on global estimation. Such results were nearly identical due to the central limit theorem effect, which further prevented the acceptance of kriging by practitioners. The second point is that a single case study using an existing property, as it was frequently performed in the past, is not sufficient because results depend not only on the tool itself, but also on the nature of the mineralization.

Jan 3, 1984

By David G. Van Hersett

The single most unique design feature of the Centralia Coal Mine is the public information program. Just what the heck does public information have to do with the development of a coal mine? The utilities of today have had to build larger and larger plants to keep down the costs of power to the consumer. The large plants require huge amounts of capital, take up large areas of real estate, consume unprecedented amounts of coal, and generate large quantities of waste products which have to be reckoned with. Coal mines supplying these plants have also increased in size and production to meet the power plants' unending appetite for coal. This means the disturbance of great areas of land and significant alterations to the landscape. Today's public is greatly concerned about the environment in which he is living. The great size of a power plant with its large stacks, coal piles, and the coal mine with its large earth movers, disturbed acreages and the resulting waste products of both operations appears to present a threat to his environment. Because of this,' many power plants and coal mines are having public acceptance problems. This has resulted in the passage of stringent legislation and in some cases brought upon construction delays for major power plants.

Jan 1, 1970

By Mark R. Bowie

Air-fall tuffs have been altered to zeolite via reaction of volcanic glass with local pore solutions at the six localities discussed in this paper. These are in the southern Basin and Range Province of Arizona and New Mexico, and in the Trans-Pecos and Coastal Plain regions of Texas. Zeolitization occurred in either closed- hydrologic, open-hydrologic, or hydrothermal groundwater systems. The tripping Spring Valley (DSV) chabazite in Arizona and the Buckhorn clinoptilolite in New Mexico originated in a late Cenozoic network of closed-hydrologic, lacustrine basins extending from southeast Arizona into southwest New Mexico. At DSV, two potentially commercial chabazite-rich tuffs are interbedded with lacustine mudstone. The main zeolite tuft contains up to 90 percent chabazite. The upper zeolite tuff averages approximately 67 precent chabazite. At the Buckhorn deposit, two clinoptilolite-rich tuffs are interbedded with lacustrine m udstone in the upper part of the Gila Conglomerate. The lower zeolite tuff contains up to 90 percent clinoptilolite. The upper zeolite tuff contains over 60 percent clinoptilolite. The Casa Piedra clinoptilolite is in the Trans-Pecos region of Texas. It formed during the Tertiary in a closed basin containing a playa lake. The Casa Piedra tuff averages about 60 percent clinoptilolite. Air-fall tuffs of the Cuchillo Negro, New Mexico and Tilden, Texas clinoptilolite deposits were zeolitically altered by either percolating ground waters in open-hydrologic systems or locally by ponded, saline, alkaline solutions. Cuchillo Negro tuffs are Tertiary and average approximately 50 percent clinoptilolite. Clinoptilolite-rich tuffs of Ebcene age have been mined from four quarries at the Tilden deposits. Tufts from the Kuykendall and Buck Martin quarries contain approximately 50 to 60 percent clinoptilolite. Low temperature, SiO2 - rich, hydrothermal solutions altered air-fall tuff to clinoptilolite and chabazite in the Foster Canyon area of the Tertiary Cedar Hills and Seldon Hills volcanic-vent zone in southcentral New Mexico. Poster Canyon tuffs contain up to 60 percent zeolite and average 40 to 50 percent clinoptilolite. The dominant mineral in the zeolite tufts at the six deposits studied is generally microcrystalline and either chabazite or clinoptilolite. Lithic fragments, unaltered volcanic glass, other zeolites, quartz, calcite, opal-CT, cristobalite, feldspar, evaporites, clays, and mafic minerals are also constituents of the tufts.

Jan 1, 1987

By Richard Keith Corbin

INTRODUCTION The purpose of this paper is to assist explorationists and entrepreneurs in deciding whether to invest in gold mining concessions in lesser developed countries (LDCs) and, if so, in deciding which LDCs provide the best opportunity. The first decision is difficult in the face of common notions that foreigners are not welcome - only their money; that the best gold deposits were worked out long ago; that even if you survive the mal- aria, and in spite of 51% control by a national partner, you manage to profit from a new discovery, they will tax your operation to death, or worse yet, nationalize it. Close scrutiny of what happened to foreign investors in LDCs during the “Mexicanizationt” of the 1930's through the "Zambianization" of the 1970's shows that many of these concerns are well-founded historically. But the trend of knowledgeable writers over the past decade seems to be away from this gloomy foreboding, and towards a cautious optimism, based on recently perceptible metamorphoses in certain LDCs' conditions and attitudes. If the explorationist or entrepreneur decides to seize on this apparent opportunity, the second decision -- which of these certain LDCs to invest in -- is doubly difficult, because of the formidable tasks of coming up with a rating system to provide rational comparison, and then gathering enough current, reliable information to apply. To assist in these tasks, a suggested rating system and case studies, respectively, follow a brief review of recent writings on the subject

Jan 1, 1987

By R. D. Grove

The control of Florida phosphate flotation plants is largely based on the visual observations of experienced operators, both in recognizing good and bad froth conditions and in making judgments about the composition of feed, concentrate and tailing streams. Chemical analyses lag the process by at least two hours and, therefore, only assist in recalibrating the operators' vision when major deviations occur. An image analysis system was developed by IMC Fertilizer to improve the accuracy and precision of the composition estimation process. It still relies on the operator taking a representative sample, but provides a measurement in about 15 minutes after that sample is obtained. Background The search for a quick and inexpensive technique for estimating phosphate flotation stream composition continues. There have been attempts at using "wet chemistry," nuclearmagnetic resonance, neutron activation, X-ray diffraction and other techniques reported on in the past several years. None of these have yet found its permanent place in the flotation plants of the Florida phosphate industry. IMC attempted to use wet chemistry to provide 7.5-minute estimates of final concentrate insol and tailings BPL. This was aborted after 36 months of plant operation due to the high operating costs for chemicals and an operator. There was also no significant economic improvement in the plant's operation. Another factor was the necessity to limit these crude analyses to only a few process streams, which were insufficient to achieve improved plant operation.

Jan 1, 1993

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998

By Ronald D. Mariani, Gregory E. Mahoski

BACKGROUND The steel and iron ore industry has had a long history of an adversarial relationship with its union and workforce. The Empire Mine during its first 40 years of existence was no exception. The "we-they" mentality went hand-in-hand with the normal by-product of an adversarial relationship, grievances and arbitration. This relationship was considered the norm and it was not until the late 1970s that consideration was given to changing this relationship. At this time the industry was involved in a fight for existence and the economic pressure demanded increasing productivity to lower cost. The challenge to survive highlighted the fact that an adversarial relationship was not conducive to attaining the needed gains in productivity. The first formal attempt to change the way we worked came in the 1980 Labor Agreement which provided for the opportunity to utilize Labor Management Participation Teams (LMPTs). LMPTs were established at the Empire Mine and they functioned for a couple of years. The teams spent most of their time working on creature-comfort issues and steered clear of dealing with productivity issues. Mine shutdowns had occurred in Michigan resulting in large layoffs, and the union was trying to add jobs and was not supporting teams that could increase productivity and eliminate jobs. When the 1986 labor negotiations began, the local union pulled their support of the LMPT process. Other initiatives were negotiated over the years to foster a change in the culture but, as with the LMPT program, they did not produce the desired culture change. It became abundantly clear: you cannot negotiate a culture change.

Jan 1, 2004

By Paul C. Miclea

With the growing urban population and the concern for traffic congestion and pollution (emissions control), public transportation is becoming more and more attractive to both city dwellers and managers. To gain access to the central area of the city, the sub- way remains the most viable alternative, despite its higher cost when compared to above ground or elevated transportation systems. There are relatively few regulations and criteria for subway ventilation, particularly when compared with mine ventilation. The main document that provides guidance and general recommendations for subway ventilation and environmental control is the Sub- way Environmental Design Handbook, published in 1976 (2nd edition) by the U.S. Department of Transportation, Office of Re- search and Development. Many of the subway systems in existence today have been designed and built with ventilation features adequate for normal train operation, but their design does not consider stringent criteria for such emergency conditions as a train fire in a tunnel. The National Fire Protection Association's Standard for Fixed Guideway Transit Systems known as NFPA 130 as well as the ASHRAE Handbook - HVAC Applications (1995) provide specific design and operation requirements for subway ventilation systems. Some of the existing, old subways are upgrading their ventilation systems to comply with the new regulations. The ability of a particular ventilation system design to provide adequate ventilation during normal and emergency conditions can be evaluated using computer modeling and simulation techniques. A train fire will cause a sudden change in the tunnel ventilation pattern by adding an unsteady and fast-growing source of heat. The hot air and gasses created by a fire will tend to flow uphill, possibly against the normal flow, producing a 'backlayering effect." To prevent this effect from happening, enough ventilation must be provided and the governing criterion to establish the required airflow is called 'Critical Velocity." Several software packages are available for special applications on tunnel and station ventilation as well as to model the spread of smoke and heat in case of a major tunnel fire, using Computational Fluid Dynamics. This paper presents the current issues in subway ventilation, as shared by the public transit community around the world, together with modern alternative design concepts vis-à-vis the more stringent emerging Fire-Life Safety criteria. Examples of recent studies and ventilation systems design are provided.

Jan 1, 1997

By M. Girardeau, J. Blondin, M. Nomine

It is well known that surface moisture and fine content have a strong influence on coal handle- ability. It is known as well that these two criteria are not sufficient to classify coals. It seems that certain types of coals are intrinsically difficult to handle, the reason being still not well understood. For these coals, moisture and fine content are aggravating factors and have to be strictly controlled in order to avoid handling problems (sticking, arching in hoppers.. .). CERCHAR has undertaken research aimed at identifying the role of intrinsic characteristics of coal with regard to handling properties. Two of them have been identified as particularly relevant : the clay content and the "colloidal particle" content. Two simple tests commonly used for soils characterisation, but never used so far for coal, have been applied to characterise a wide range of coals. The derived classification has been compared with an empirical classification based on dry screening tests. DRY SCREENING EXPERIMENTS Dry screening at a relatively low screen aperture (e.g. 6 mm) may be considered as a good tool for assessing the handleability of wet fine coals. Difficult to handle coals always lead to rapid plugging of the screen whilst easy to handle coals can be screened without difficulties. A wide range of coals has been selected with various ash, moisture and fine content. Their characteristics are given in table 1. The experimental set-up consists of a 2 t.p.h vibrating screen equipped with two automatic re- cording balances collecting respectively oversize and undersize material. Provided that the feeding rate is accurately controlled, good reproducibility in the percentage passing as a function of time is obtained. Figure 1 gives examples of different types of relations that were obtained. - LMC coal exhibits a behaviour that is characteristic of easy to handle coals : % passing remains constant in the course of the screening operation. - B coal behaves slightly differently. One observe first a rapid decrease in the flow passing through the screen and then a stabilization. Hand- ling properties can be considered as intermediately - The curve relative to LHB coal is characteristic of a very difficult to handle coal that blinds the screen almost immediately. It is necessary to point out the fact that the classification derived from screening experiments is a global one, including not only intrinsic factors but also external factors such as fines and moisture content. The role of surface moisture is in particular very important and variations in this parameter may upset the empirical classification

Jan 1, 1988

By V. F. Hollister

The Ivanhoe district, Elko Co., Nevada, is located in the northern part of the Carlin trend of epithermal gold-deposits, but it differs from most mines in the belt in that Ivanhoe epithermal gold ores are found in 14 m.y. old Tertiary volcanic rocks as well as in the better known Paleozoic sediments. Within the Ivanhoe district U.S. Steel (USX Corp.) discovered and partly developed an epithermal gold deposit largely hosted by Tertiary rocks that, with a 167-hole program, was found to contain an open ended probable reserve of 4.0 million tons of heap leachable open pit oxide ore averaging .042 opt Au. An additional 4.0 m.t. of open pit sulfide ore averaging .048 opt Au were also developed. The U.S. Steel orebody was a blind discovery in that it just barely outcropped, and it was found by drilling through a premineral but barren cover of altered Tertiary volcanic rocks.

Jan 1, 1988

By A. W. Easton

INTRODUCTION This presentation will address changing trade patterns in lead and zinc from the mid 1980's to the early 1990's. Lead and zinc deposits are developed where they are found in many parts of the world. As many are not in major consuming areas, there is significant international trade in both lead and zinc concentrates and in refined metal. Forecasts of mine production have been analyzed, primary and secondary refining capacity and future consumption levels have been estimated, to arrive at the changes in trade patterns expected during the next half decade. This paper will first address concentrate and metal trade patterns for lead, then for zinc, and finally some remarks will be made on captive mine/smelter agreements, on regional trade agreements, and on barriers to trade. I would like to note that. there is only a small international trade in semi-fabricated products such as zinc die-cast alloy, zinc oxide, lead oxide or rolled lead and zinc. Most of the world's zinc and lead trade is in concentrates or refined metal. LEAD TRADE PATTERNS Lead Concentrates Six countries are exporters of lead concentrates. Three countries are in the Americas - Canada, Mexico and Peru, with two in Africa - Morocco and South Africa (including Namibia), and Australia. Due to the limited growth in lead consumption and consequent minimum investment in new production facilities expected, mine and metal production levels and future consumption levels are not forecast to change much between now and 1990. Accordingly, the tonnage of lead concentrates exported will remain essentially unchanged from these six countries. In total they have exports of 480,000 metric tonnes (metal content)* and by 1990 this figure is forecast to rise only slightly with the only increase taking place in Mexico. The concentrates are exported mainly to Western Europe & Japan. Apart from these six main lead concentrate exporting nations, a significant change is expected in the United States. In the mid * All concentrate trade figures are based on metal content

Jan 1, 1986

By Cary Hirner, Tony Cicchetti, Daniel G. Cressman, Jeff F. Wallace, Prapan Dave, Malcolm Sheehan

The City of Toronto is currently undertaking the design of the Black Creek Sanitary Trunk Sewer Relief Project. The Project involves the design and construction of approximately 20 kilometers of tunnel through shale rock of the Georgian Bay Formation and soft ground soil consisting of glacial till, glaciolacustrine and glaciofluvial sand, silt and clay deposits. The 20 kilometer alignment has been split into four sections of tunnel (1) the 12.9 km, 3.0 meter diameter Keele Trunk Relief Sewer, (2) the 2.9 km, 1.5 meter diameter Keele Trunk Relief Sewer Microtunnel, (3) the 3.9 km, 1.8 meter diameter Jane Relief Sewer, and (4) the 0.5 km,1.5 meter diameter Combined Sewer Overflow Storage Tunnel. Additionally, one 40,000 m3 wet weather flow (WWF) storage tank and numerous WWF connection shafts are to be constructed and connected to the Keele Trunk Relief Sewer to pick up WWF from existing outfalls currently discharging to Black Creek. The anticipated cost of the project is approximately $600 million with construction of the Project’s Phase One Works, the Keele Trunk Relief Sewer and Jane Relief Sewer, scheduled to commence in 2025. This paper provides an overview of the project in terms of the anticipated geotechnical conditions, the scheme developed for design and construction of the tunnel system and the associated schedule and procurement strategy. The consideration for planning of the tunnel system, the risk-based analysis and decision-making process are presented and form the basis for selection of a precast tunnel lining support method in the rock and soft ground.

Jun 13, 2023

By Gerald Zink

The work described in this paper is sponsored by the Bureau of Mines for the purpose of accelerating the drilling of roof bolt holes and reducing the quantity of materials needed to fill each hole. A high-speed small-hole drill has been fabricated and tested in the laboratory. High priority has been given to safety, reliability, and compatibility with underground coal mining practice. The combination of drag bits and erosion jets is designed for high penetration rate, long wear life, and uniform hole size in a variety of sedimentary strata. A reliable high-pressure high-speed swivel feeds water to the bit at about 4 1/min (1 gpm). The duplex hydraulic intensifier which drives the jets has a 40 kw capacity (30hhp) Results of recent full-scale laboratory testing are presented.

Jan 1, 1978

By -Ing. Alexander Dohmen, -Ing. Gunter Hagen

In order to improve the thermal environment in deep coal mines, two principles can be applied: firstly, the reduction of heat flow into the air stream, and secondly, cooling of the air current. First of all one should try to minimize the heat-flow rate into the air stream underground by the mining method itself. German coal mines prefer the caving system of advance longwall mining simultaneously on several levels with ascensional ventilation. The effects of descensional ventilation and different face ventilation layouts on the thermal environment are discussed, besides the effects of pneumatic and hydraulic stowing. Application of all mentioned possibilities together instead of today's common mining method would reduce the Basis Effective Temperature considerably. In the future, hydraulic haulage of the raw coal underground substituting belt conveying may be a mean to improve the thermal environment further on. If the working depth of a coal mine exceeds about 1 000 m, cooling of the air stream under-ground will be necessary in addition to the above- mentioned measures. Generally, in German coal mines different types of indirect-contact air coolers located nearby the face are used. Besides such so-called cooling coils, large-capacity spray coolers are used because they work more efficiently if a great cooling duty is demanded. Common designs of both air cooler types and their possible utilizations are demonstrated. Deep coal mines in the Federal Republic of Germany need up to 30 MW of total cooling capacity. Mostly, the needed chilled water is cooled by large-capacity refrigeration plants placed on surface, underground or in combination of the above locations nearby the shaft. Due to the increasing underground cooling demand, the distribution of the chilled water and the control of the air coolers become more important. The conception of a control system is discussed, which is based on the control of the single air coolers. All air coolers of a mine are controlled so as to optimize the supply water-flow rate with regard to the demanded cooling duty of the coolers. On the next upper level of the control system the whole chilled water network and finally the refrigeration plant is regulated. Present research work deals with the use of ice for cooling mines. There are some unsolved problems in this new cooling concept. Some aspects of making, transporting and storing ice are described. The use of ice may probably reduce the cooling costs of German collieries, which are at present approximately 4 DM per ton of saleable coal.

Jan 1, 1989

By Zhou Aimin, Xie Benxian, Song Chongkai, He Zhexiang, Zhang Changqing

Concrete pump is commonly used to transport paste and high-density backfill over relatively long horizontal distances. However, since the investment in concrete-pump and its operating cost are quite high, the application of paste backfill is limited in China. For seeking an alternative paste backfill technology, squeezed-transport technique and equipment has been developed. The new technology has been adopted at Tonglushan Copper Mine of Daye Nonferrous Metals Corp for overcoming a difficulty of long-distance transportation of high-density cemented tailings fill. It proved to be an alternative paste backfill method with simple backfill system, saving investment and low operating costs.

Jan 1, 2001

By Jakub Nalaskowski, Robert Aranowski, Jan D. Miller, Jan Hupka

The polyolefins, mainly polyethylene and polypropylene particles, are used as an active filler material in polymer-polymer blends and are being used for the modification of asphalts to pro- vide for high-load, low-maintenance road construction. The stability of prepared blends depends on the interaction forces between polyolefinfiller and the matrix and/or between two polyolefin filler surfaces in the polymeric matrix. Various additives, especially surfactants, can be used to modify these interactions and improve the stability of polymer blends. In current research, direct force measurements between polyethylene (PE) surfaces in n-tetradecane, a nonaqueous solvent imitating the polymer matrix, were performed in the presence of lauric acid and dodecylamine using the AFM colloidal probe technique. Additionally, the zeta potential of polyethylene particles in the presence of surfactants in n-tetradecane was measured. The interac- tion forces between the polyethylene surfaces can be described as a combination of electrical double layer repulsion and steric repulsion between polymer chains extending into the n-tetradecane sol- vent. The addition of surfactant to the n-tetradecane was found to increase the repulsive force between the PE surfaces. The strongest repulsion was observed for dodecylamine. Additionally, the temperature efect on the interactions in n-tetradecane was studied. It was found that steric repulsion between the PE surfaces increases with an increase in temperature.

Jan 1, 2003

By M. Javier

Mining is most vital to our civilization. The current model of extraction technologies was designed without an expiration date and therefore generates environmental effects that are severe. Mining results-effects have been traditionally accepted and promoted around the world since Agricola described mining. Mining’s extractive technology is outmoded and standardized and inflicts negative results-effects on nature. Those results are liabilities dimensioned in legacy and type of nature able to provide to the next generation. In order to design, build, and transform mining effectively we must ask: What is the vision of mining? Mining has exhausted its antiquated mindset of over 200 years. Now it merely pilots us into crisis. The challenge today is to resolve the socio-environmental conflicts globally with a radical transformation of the mining business at its core. This paper illustrates that old mindsets must be radically updated or the concept of mining will continue to be trapped. Changing the mindset will enable discourse toward a new destiny. These changes will change mining for the better. We are able to evolve and get different results than in the past by pursuing a new legacy to live in harmony with nature for many more years. INTRODUCTION Minerals are supported by the law of conservation of mass. They have the unique characteristics of being finite, exhaustible, unsustainable, and have an unlimited recycling. Minerals are also hard to find because we have explored and virtually exhausted the whole crust of this planet. Orebodies that used to be considered world class are almost exhausted. In the short-term, it would be almost impossible to find new world class orebodies unless they could be extracted from the ocean, the moon or other planets like Mars. Mining is designed to extract minerals for profit and to keep civilizations functioning. Subsequently, mining has a social issue that is an economic issue. Thus, mining must be seen as a real and urgent subject matter that must be protected in thought and in fact. The actual mining extractive technologies produce disequilibrium in nature that has unknowable effects in the course of time. History is filled with the negative consequences to nature due to the model of mining. The evidence of these effects from mining’s model is apparent in repetitive liabilities around the world. The effects negatively impact nature without any consideration to how these impacts will manifest themselves in the future. This model of mining is widespread and extolled in every corner of the planet because it is old and familiar and although wasteful and outdated, we do not want to expend financial resources from profit or elsewhere to explore newer ideas, methods, or technologies that would help us keep a sustainable harmony between extraction and nature. We need a new vision for mining today and to explore in what way academia can and must play a social role in the teaching of mining methods. We need to question why we continue to support the old model of mining and to ask ourselves what prevents us from updating and modernizing the method of mining to meet the current needs not just of civilization but the urgent needs to undo some of the damage to the earth. This paper hopes to develop a method to change our minds about how we conduct the business of mining in order to control the destiny of it.

Jan 1, 2019

By Jacques H. Nantel, James F. Archibald

INTRODUCTION The use of backfill materials in underground mines has been practiced for many years and the trend indicates an increased utilization of backfill materials in the future. The authors believe that underground uranium mines will follow this trend. Classified mill tailings, approximately 65 to 70 percent solids by weight, are the most common fill materials. Cement or other pozzolanic agents can be added to the fill materials to improve their strength characteristics. Dry fill, in the form of development muck, can also be introduced in the stopes to fill the voids left by the action of mining the ore. This paper deals mainly with the assessment of the radiation properties of backfill materials originating from uranium ores processed in the milling facilities of uranium mines. The paper outlines the results of studies conducted in the laboratory and in the field on radon emanating characteristics of backfill materials utilized in uranium mines. ROLE OF BACKFILL IN UNDERGROUND MINES Fill is a structural component of any underground mining system. In general terms, the purpose of filling is to provide ground support, to make recovery of pillars possible to provide a working platform for men and equipment, and to allow for waste disposal. There are other inherent advantages to the use of backfill underground, the most important being to: a) permit mining at greater depth b) allow the adoption of highly mechanized mining techniques c) reduce dilution from wall sloughing d) minimize roof and wall failures e) allow the mining of irregular orebodies, and f) to provide better control of ventilation. The last item is of great importance to uranium mine operators. The introduction of radioactive backfill underground represents an additional source of radioactive contamination; on the other hand the introduction of backfill in the mine reduces the volumes of workings to be ventilated. As a consequence the general mine ventilation is more effective and in most cases, better air quality can be maintained in the mine. There are of course some disadvantages associated with the use of hydraulic fills in underground mines, notably: a) the handling of large additional quantities of water from the underground workings b) delays caused by the cyclical nature of the backfilling operations c) additional costs associated with stope preparation and placement of backfill in stopes, and d) "spills" caused by bursting pipes or fill escaping from fill retaining structures. ASSESSMENT OF NON-RADIOACTIVE PROPERTIES Although the radioactive properties of mine backfills are discussed in this paper many other physical and chemical properties of backfills must also be considered. These properties include: a) percolation rate and permeability parameters b) viscosity versus pulp density relationships c) size analyses d) settling times e) settled bulk density f) solids specific gravity g) void ratio and porosity h) optimal percentage solids by weight i) specific gravity of slurry j) addition of cement:cement retention and loss k) unconfined compressive strength 1) triaxial compressive strength m) shear strength (non-cemented)

Jan 1, 1981