Search Documents

By Michael J. Martinson

Mine ventilation has been evolving as a separate field of study since the middle of the nineteenth century, and some practitioners might assert that mine ventilation has come of age as a discrete field in the last decade or two. Certainly the extensive literature now available on many facets of the subject, international congresses, the phenomenal upsurge of interest in mine ventilation in the United States of America since the passage of the 1969 Federal Coal Mine Health and Safety Act, and the existence of a virile mine ventilation society in South Africa with its own monthly journal all point in this general direction; on the other hand the lack of a rational design philosophy for ventilating mines suggests that mine ventilation may still be on the threshold of maturity. To generalise, many mines are still ventilated simply to comply with arbitrary statutory regulations, or with some ill- defined notion of reducing the risk of injury associated with the atmospheric environment in mines. Hazards in this category include oxygen deficiency, toxic gases and particulates, pathogenic gases and aerosols, explosive gases and dusts, and extreme climatic conditions. Although many mines are ventilated to abate the risk of a specific injury, there is inevitably a spectrum of latent and patent atmospheric hazards in every underground mine. The writer (Martinson, 1976) has previously drawn attention to the dearth of material on mine ventilation design in textbooks and technical journals, and to the overly simplistic approach sometimes adopted in the limited work published on the subject. The available material undoubtedly contains useful technical data and persuasive views, but seemingly could not be used to demonstrate the rational engineering approach to mine ventilation design advocated in the present paper. The range of papers presented at the 1975 and 1979 Congresses, and the diverse interests and affiliations of delegates, emphasize the multidisciplinary nature of mine ventilation. Easily recognisable topics include: Fluid mechanics Thermodynamics Heat transfer Flames and combustion Aerosol physics Radiation physics Geophysics and geomechanics Occupational diseases Epidemiology Industrial physiology Despite these heterogeneous connections there can be little doubt that mine ventilation is fundamentally an engineering field and not an off-shoot of some other branch of applied science. The distinction is important because mine ventilation has essentially developed as a form of ad hoc trouble- shooting, and in the process mining engineers have relied extensively on the assistance of experts drawn from other fields. Thus historically there has been limited scope for the application of rational design techniques, and personnel enlisted from scientific disciplines may not have had any previous experience in the intricacies of engineering design.

Jan 1, 1980

By C. O. Brawner

Geotechnical theory is reasonably well developed to evaluate the stability of rock slopes for surface mining projects. It is recognized widely that structural geology, groundwater pressures and seepage, and blasting are key factors that influence the stability of such slopes. However, our capabilities to model the structural geology, to determine rock strength along discontinuities and to model groundwater pressures and profiles are still very approximate. Plus, practical experience of the rock mechanics engineer is one of the most important inputs to the design of open pit mine slopes. To emphasize the use of practical experience, a number of case examples are reviewed here to illustrate key factors that must be evaluated. These case examples comprise a continuation of other examples presented in the Jackling lecture to the SME in 1986 by the author (Brawner, 1986). Example no. 1 In situations where the orientation of geologic discontinuities will not influence rock slope stability, extremely steep slopes can be developed, providing blasting and groundwater are controlled adequately. An excellent example of this condition is the Republic Pit of Cleveland Cliffs in Michigan. The rock slope in Fig. 1 isover400fthigh with a face angle of approximately 80". The rock is hard and massive, and blasting was carefully controlled. The original design of this slope was at45", a standard design angle at that time. The competency of the rock and the lack of critical geologic structure subsequently was recognized, and the much steeper slope angle was developed successfully. The rock proved to be so competent and the slope sufficiently stable that an underground crusher station was developed with entrance and exit tunnels excavated at the toe of the steep wall.

Jan 1, 1991

By Ezra L. Kotzin

Foundry sands used to make molds and cores have been very important to foundrymen since metal casting began hundreds of years ago. They are now used in two basic ways, either in an uncured green state, grain-bonded with clay, or in a set cured state achieved with sand that, mixed with resins or oils, is cured by baking or by chemical reaction. The term sand, as generally used in foundry applications, may best be defined as material composed of granular particles of mineral matter, ranging in size from 0.5 to 2 mm in diameter. The particulate nature of the material may be the result of natural disaggregation or result from crushing and screening of rock or ceramic materials; crushed and screened products are sometimes referred to as "manufactured sand." Based on their origin, the raw materials for foundry sands will vary in grain shape, grain composition, relative surface, grain size, and grain distribution patterns. These properties, in addition to their chemical analyses, sintering point and expansion characteristics play an important part in the choice of sands used as the base molding or core aggregate in metal casting. Sands with different origins may be blended to produce specific compositions and grain size distributions. Early raw material requirements based on silica sand were simple, but as sand/core/mold technology became more sophisticated and binder technology continued to be developed, the physical properties of foundry sands became more critical. Silica sand is composed of quartz, along with, in most cases, small amounts of feldspar, mica, clay, and other common minerals. Although silica sand is still by far the most widely used base material in production of molds and cores used in metal casting, other natural mineral sands have their own unique characteristics and fill an important niche in special applications. As a result, the term sand, as applied to raw materials used in metal casting, has been logically extended to include granular materials composed of a group of minerals other than quartz. With increased use in advanced foundry technology, natural sands composed predominantly of zircon, chromite, staurolite, or aluminum silicate minerals and of granularized chromite or olivine, have become more than just alternate materials and are now commonly classified as sands when the grains are sand-sized. Foundry sand may therefore be considered under two categories: silica sands and non-silica sands.

Jan 1, 1994

By Timothy J. L. Haley

The population of the Rhondda underwent a spectacular rise as energy requirements increased in the mid-nineteenth century. The rapid expansion in the demand for coal transformed the population from a few hundred sheep farmers in the eighteenth century, into a thriving coal-based economy of some 168,000 inhabitants in 1924. Cochrane (1958) put forward the idea that the differential migration of more healthy men from the Rhondda Fach during the period since the peak of population in 1924, had resulted in a respiratorily-impaired residual population. This idea was used to explain the reduced survival of miners & ex-miners aged 55-64 years from this valley as compared with a similar group from Leigh in England. Using the proportional hazards regression model of Cox (1972), a comparison is presented of the exact twenty-year mortality of migrant and non-migrant miners & ex-miners and non-miners from the Rhondda Fach. The analysis makes allowance for the pattern of migration throughout the twenty-year period and adjusts for the confounding influence of variables such as age, occupation, and category of Simple Pneumoconiosis. The results indicate that contrary to expectation, any effect of migration upon mortality is manifest only at higher ages. Furthermore, for these men, the migrants have worse survival than the non-migrants. Also, contrary to expectation in the light of work by the British Coal Board (Soutar et al, 1982) there is no evidence of increased mortality amongst men with category I simple pneumoconiosis. INTRODUCTION [The Background] In September 1950, the Medical Research Council's Pneumoconiosis Research Unit at Penarth in South Wales initiated a study of lung disease in the smaller of the two Rhondda Valleys - the Rhondda Fach. At that time, the primary interest was in assessing the possible role of pulmonary tuberculosis in the aetiology of the serious (complicated) form of coal-workers' pneumoconiosis, i.e. Progressive Massive Fibrosis. The Rhondda Fach is situated in South Wales, some 25 km north-west of Cardiff. It is a narrow winding river valley about 12 km long, and approximately 1 km wide. The name means Little Rhondda; there is a corresponding Big Rhondda - the Rhondda Fawr. Together, these valleys make-up most of the Rhondda Municipal Borough. Following the initial survey in 1950-51, there have been several follow-up studies of this population. In addition, amongst numerous other investigations to unravel the aetiology of coal-workers' pneumoconiosis, Medical Research Council teams studied various other (sample) population groups, particularly in the industrial areas of the adjacent Aberdare Valley, and in Staveley and Leigh in England. The agricultural areas of the Vale of Glamorgan in South Wales, and Annandale in Scotland were also studied.

Jan 1, 1993

By Owen Kingman

This paper discusses grade control of ore mined at Tennessee Copper Company. Mineral zoning and the influence of mineral zoning on production grade are discussed and illustrated. In order to make this most understandable we will first briefly examine the flow sheet at Tennessee Copper to indicate the path followed by the raw material toward its various products. Then we will discuss some of the characteristics of the ore, specifically the distribution or zoning of the minerals in the ore which cause some of the grade control problems. Finally we Will discuss same of the steps that are taken to combat these problems. The flow sheet is illustrated as plate 1. The mined ore is allied to obtain 3 usable products: Copper concentrate, iron concentrate and zinc concentrate. Copper is found in the mineral chalcopyrite, CuFeS2. The Copper concentrate from the mill contains about 20% copper. These concentrates are smelter in reverberatory furnace and Pierce-Smith converter to blister copper containing 99.2% copper. Some metal is cast into 360 lb. pigs and shipped to the refinery. Some is made into shot copper and used locally to make copper sulfate. Copper sulfate is marketed from Copperhill principally in fungicides.

Jan 1, 1960

By L. A. Weakly

The requirement for large diesel equipment to be certified for operation in potential gassy mine conditions in western oil shale mines motivated a design engineering project to develop a dry exhaust conditioner system to replace traditional wet scrubber technology. A dry exhaust conditioner for a 600 hp diesel engine was designed and fabricated for retrofitting into a 600 hp underground haul truck. The engine conditioner package was taken through a simulated certification process at the Mine Safety and Health Administration (MSHA) Approval and Certification Center to refine test procedures and to uncover any flaws in the dry exhaust conditioner technology. The engine/exhaust system was then installed in an underground haul truck to demonstrate mine-worthiness of the system. No inherent features were discovered that would prevent the system from meeting test requirements. However, the tests did point out several deficiencies in the prototype design.

Jan 1, 1989

By Dan Xu, Song. Gaofeng, Zhen Wei, Zhihe Wang

"With decades of development and exploitation of coal resources, the first batch of coal-based cities in China are faced with the challenge of economic shrinkage and environment deterioration due to the rapid resource exhaustion and single industrial structure. Nowadays, as a policy instrument for sustainable development, a nationwide circular economy mode has been implemented during the design and planning of new coal mines to address the challenges arising from economic depression, energy shortage and environmental pollution. It should be mentioned that several national coal bases and most of constructing and planning coal mines are located in the underdeveloped areas in the Midwest of China which means that the distance between supply source for energy resources and the demand market is significantly increased. Moreover, the eco-environment vulnerability in these rural areas should be taken into serious consideration. Therefore, after comparison and analysis with several successful coals based and diversified circular industry chains, this paper presents a case study on the circular economy park (CEP) in the northwest of Shanxi Province. Through discussion and evaluation of ‘1+3’ mode in the CEP, this paper aims to enhance the understanding of coal mine sustainable development and shed light on the relationship between resources exploitation and local economic development.1. Introduction Earth's fossil energy resources are crucial to both economic prosperity and human well-being. Taking coal mining and consumption as an example, coal has been widely used to generate electricity since the 1880s, and more than 40% of the electricity globally is provided by coal nowadays. However, as one of the largest worldwide anthropogenic sources of carbon dioxide releases (39% of global CO2 emissions), the extensive excavation and utilization of coal have brought a series of side effects to both natural environment and human health (Wikipedia 2017). As the largest coal producer and consumer in the world, China is leading the list over the past decade (as shown in Figure 1) (BP 2016). The excavation and utilization of coal resources have made a crucial impact and irreplaceable contributions to China’s industrial development and economic prosperity (Long and Zhang 2009). However, most of the air pollution in China are also attributed to the simplified and extensive exploitation of coal resources, e.g., 90% of SO2, 67% of NO, 70% of smoke ashes, and 70% of CO2 in 2016. In recent years, most of North China and the developed coastal areas were wreathed by smog due largely to the oversimplified extensive utilization of coal, suggesting that the development patterns of coal utilization need to be reshaped so that both the biosphere and humanity could thrive equally."

Jan 1, 2017

By E. P. Chen

This investigation involves the development of a two-dimensional continuum model for the description of jointed media. Specific application of the model developed here is for the potential geological disposal of radioactive waste in the layered sequence of fractured volcanic tuffs under Yucca Mountain, Nevada. This is currently being considered by the Nevada Nuclear Waste Storage Investigations Project, administered by the Nevada Operations Office of the U. S. Department of Energy. The work is a direct extension of the previous constitutive model by Thomas (1982) based on the work of Morland (1974). The main feature of Thomas's model includes two parts: a continuum approximation based on average discontinuous displacements across jointing planes within a representative elementary volume, and a material constitutive description based on linear elastic matrix material behavior and nonlinear normal and shear joint behavior between jointing planes. The mathematical descriptions by Morland based on strain partitioning among matrix material and joint sets allow multiple joint sets to be included in the model. However, the initial effort by Thomas to implement Morland's mathematical model included only a single set of joints. To make the material model reflect more realistically the field characteristics, the current work extended Thomas's model to include orthogonal sets of joints. The nonlinear elastic normal joint behavior is retained in the present model. Joint shear response was treated as linear elastic in the shear stress versus shear displacement relationship prior to attaining the stress level governed by the Coulomb friction criterion. Beyond this critical stress value, a horizontal line is assumed in the work of Thomas for the shear stress versus slip displacement relationship such that the shear stress remains constant while the slip displacement increases. Motivated by experimental observation, the current work generalized the shear response to allow a line with finite slope to govern the slip behavior beyond the critical stress level. The inclusion of an orthogonal joint set led to a fourth order algebraic equation for the calculation of stress increments as compared with a second order equation obtained by Thomas. The nonlinear elastic normal joint behavior provides criteria which allow the selection of the root that yields result consistent with the loading condition of the problem. The addition of one more set of joints provides the potential to more completely study the interaction of various parameters representing the characteristics of jointed media behavior. Material responses to various loading conditions based on the present constitutive model have been obtained. These results and their comparisons to single joint set results are presented. Formulation of the Continuum Model The inclusion of an orthogonal joint set led to a fourth order algebraic equation for the calculation of stress increments as compared with a second order equation obtained by Thomas. The nonlinear elastic normal joint behavior provides criteria which allow the selection of the root that yields result consistent with the loading condition of the problem. The addition of one more set of joints provides the potential to more completely study the interaction of various parameters representing the characteristics of jointed media behavior. Material responses to various loading conditions based on the present constitutive model have been obtained. These results and their comparisons to single joint set results are presented. Let us focus our attention on the set with normal n first. Let the jump dilation displacement and jump slip dasplacement associated with the [ ]

Jan 1, 1986

By Ralph J. Holmes

INTRODUCTION Although sampling techniques used by the mineral industry continue to improve, sampling is still an area which is often neglected. Frequently, sampling and sample preparation requirements are left to personnel who do not fully appreciate the significance and importance of sampling. Consequently, plant evaluations are often based on seriously biased samples, leading to mass balance problems and incorrect assessments of plant performance. The basic rule for correct sampling is that each particle of ore or concentrate must have an equal probability of being collected and becoming part of the final sample for analysis. If this is not the case, bias is easily introduced and the sample is not representative. For example, when ore is travelling on a conveyor belt, the lumps may come to the surface. Consequently, a grab sample taken from the top layers only will contain a greater proportion of lumps, i.e., the sample is biased. There seems little point in making major investments in precise analytical equipment, if the sample provided to the analytical laboratory is not representative in the first place. This paper describes the sources of plant sampling errors, reviews the methods of estimating their variance, presents some worked examples, and provides some practical advice for reducing sampling errors and their variance. SAMPLING BASICS It is generally accepted that the discharge point of an ore or concentrate stream is the most suitable sampling location. The stream can be intersected at regular intervals, and representative samples, referred to as increments, can be obtained by taking a complete cross-section of the ore stream. On the other hand, sampling devices that take part of the stream on an intermittent or continuous basis, e.g., a bleed from a pipe, may introduce serious bias, and should be avoided at all cost. Increments may be taken at completely random times or tonnages during the sampling operation, in which case it is called "random sampling". However, as shown below, random sampling results in the largest sampling variance, so it is preferable to divide the ore or concentrate stream up into strata of equal time or mass and take one increment from each stratum. This operation is called stratified sampling and may be either systematic sampling, in which increments are taken at the same point in each stratum, or stratified random sampling in which the increments are taken at random within each stratum. Systematic sampling results in the smallest sampling variance, provided periodic variations in quality or quantity are not present which may coincide with, or approximate to, any multiples of the proposed sampling interval. In such cases, it is strongly recommended that stratified random sampling within fixed time or mass intervals be carried out.

Jan 1, 1991

By Mark D. Cocker

Georgia?s Coastal Plain contains important sediment-hosted deposits of kaolin, bauxite, heavy minerals, sand and gravel, silica sand, carbonate rocks, and fuller's earth. Genesis, location, size, and quality of these deposits are related to terrestrial, marginal-marine, and shallow-marine depositional environments that evolved during development of the southeastern North American continental margin. From the Cretaceous to Eocene, kaolin and aluminous sediments, which contained mica and feldspar, were deposited in marshes, ponds, and bays located in deltas or behind barrier islands or spits. In the Cretaceous sediments, subsequent, intensive, in situ chemical weathering converted mica and feldspar to kaolin. Kaolin eroded from deeply weathered granitic rocks in the Georgia Piedmont was deposited in Paleocene to Eocene sediments. Subaerial exposure during the Paleocene and Eocene desilicated kaolin sediments to form bauxite. Sand and gravel, silica sand, and heavy-mineral deposits are found in Cretaceous to Eocene fluvial and shoreline sediments, some of which are associated with the kaolin and bauxite deposits. Source areas, intensive chemical weathering in the source areas, transportation by the major river systems, and deposition of these clastic sediments in similar or nearby depositional environments are common elements in the genesis of these resources. Eocene fuller's-earth deposits are opaline claystones in near-shore to shelf-marine sediments, which overlie the kaolin-bearing strata. These claystones consist of montmorillonite and opal/ cristobalite derived from alteration of diatom frustules. Favorable ocean currents provided nutrients for extensive planktonic growth. Oligocene and Miocene carbonate rocks were deposited principally in shallow-marine embayments. Miocene fuller's- earth deposits in south-central Georgia and adjacent parts of Florida contain palygorskite that formed through alteration of montmorillonite/bentonite clays. Deposition and subsequent alteration occurred in shallow, brackish-water lagoons under relatively high temperature and subtropical conditions. Middle and late Miocene phosphorites were deposited on a shallow carbonate-shelf platform and were concentrated proximal to anticlinal highs. Numerous marine transgressions and regressions since the beginning of the Pleistocene resulted in shoreline detrital deposits containing aggregate sand, silica sand, and heavy minerals on the Coastal Plain and on the Continental Shelf. Major river systems transported sand and heavy minerals from source areas in the Piedmont and Upper Coastal Plain. Deltaic deposits of heavy minerals were redistributed and concentrated by longshore currents, shoreline processes, and aeolian processes. River-terrace sediments, correlative with the Pleistocene shoreline deposits, contain sand-and-gravel and silica-sand deposits. Aeolian dune sands are associated with both Pleistocene fluvial and shoreline detrital deposits.

Jan 1, 1999

By Xingtian Hui

Due to variation in the geological natures of underground coal mines, especially when the entry?s roof/floor are soft and weak, it is vital to select an appropriate roof bolting system to ensure entry stability, coal production, and miner?s safety. No single type of roof bolting system perfectly fits for various types of openings, even within a mine. From the results of continuous research on roof bolting in the past 10 years, the authors used a novel concept of ?Self-screwed? mechanism to develop five types of roof bolting under a patented Self-screwed Tube Bolting System (STBS)©. This roof tube bolting system can be applied to meet different timing-strength requirements under varying mining conditions to increases the stability of mine workings. This paper outlines the structures and mechanisms of the five types of self-screwed roof bolting systems along with their unique characteristics, advantages and disadvantages, and limitations. Two case studies using a Self-screwed Tube Drilling Bolting (STDB)©, one of the five types of roof tube bolting systems, show the principles of mechanics, suitability and parameters for a mine entry and a subway?s construction/ supports in the surrounding soil layers and soft sands, respectively are presented. The results confirmed the time dependent anchorage capacity between predictions and on-site measurements. Currently, other types of the self-screwed tube bolting systems are also being applied to different coal mines in China to increase entry stability and meet safety requirements.

Jan 1, 2009

By Malcolm J. A. Swallow

In June 1997, Mount Polley Mining Corp., a division of Imperial Metals Corp., completed an 18-kt/d (20,000 stpd) gold-copper mine in central British Columbia. Fast-track construction and extensive off-site construction were combined with an aggressive project schedule to build the mine in one year. This was three months ahead of schedule and 7% below the original budgeted cost of C$123.5 million (US$86 million). The Mount Polley Mine was the first of a crop of new open-pit mines in British Columbia. The Mount Polley deposit has been explored since the mid-1960s when it was known as the Cariboo Bell property. From 1988, Imperial Metals completed drilling and metallurgical testing on the deposit. This led to a positive feasibility study and the issuing of a Mine Development Certificate in 1991. The project was later shelved due to falling metals prices and the strength of the Canadian dollar. In late 1994, Imperial merged with Bethlehem Resources Corp., bringing a strong project team to the company. The deposit was then re-evaluated, ore reserves revised and additional metallurgical testing was performed. The result was a revised project plan. It indi¬cated that a financially feasible gold-copper mining op¬eration could be constructed at Mount Polley operat¬ing at 18 kt/d (20,000 stpd). In May 1996, an agreement was reached with Sumitomo to jointly fund and develop the project. By early 1996, final plans and permits were being prepared for construction, scheduled to begin in June 1.996. On-site work commenced on May 28th, 1996 following lifting of the road restrictions. Mount Polley Mining acted as its own project manager. Construction was completed in June 1997. Plant runup took place in July and full production began in August. Through the use of off-site construction and the careful selection of contractors who cooperated to complete some 12 major contracts, it was possible to reduce the mine and mill construction time from 16 to 13 months. Off-site construction for all possible aspects of the work and the use of used milling equipment allowed significant savings in overall capital costs of construction. Total costs for the mine, mill and tailings dam equated to C$6,100/t/d of throughput. The mill design incorporates several innovative features that will aid maintenance and allow flexibility of operation. The operation is designed as a closed system and meets the latest environmental standards. Concentrate is shipped by truck to the port of Vancouver. Revenues for the mine come from the mine's production of about 3.1 t/a (100,000 oz/year) of gold and 10.8 kt/a (24 million lbs/year) of copper during the first four years of the 12-year mine life.

Jan 1, 1998

By R. V. Huff

An in-situ leaching technology was developed for copper oxide minerals at the Van Dyke deposit at Miami, Arizona. This paper discusses the process developed and the activities that occurred. Specifically, the activities of well installation, deposit data acquisition, leach testing and equipment development are reviewed. Installation and operation of the test facility are described.

Jan 1, 1988

By C. A. Rowland

Introduction Comminution efficiency, as measured by relating circuit performance to the consumption of energy and the time and money required, can be determined by the following comparisons: Theoretical efficiency: The theoretical efficiency is the actual energy consumed in a comminution machine compared to the theoretical energy required to perform the size reduction. Operating efficiency: The operating efficiency compares the operating work index of a comminution machine to the Bond work index of the ore, as determined from bench-scale crushing and grindability tests or pilot-plant tests. Economic efficiency: The economic efficiency is comparing income from production to predicted income. Key words: Comminution, Bond work index, Grinding, Crushing, Ball mills, Rod Mills. Theoretical efficiency In their book Process Engineering of Size Reduction - Ball Milling, Austin, Klimpel and Luckie (1984) stated that, "We have tried throughout this book to avoid the use of the term grinding efficiency because the degree of conversion of energy to useful size reduction is an ill-defined concept. In addition, the only efficiency, which is of real interest, is the meeting of product specifications at the minimum cost." The National Materials Advisory Board, National Re¬search Council (1981), in the abstract of a report titled "Comminution and Energy Consumption" stated that, "Up to 99% of the energy consumed in grinding these ores may be expended in the movement of machinery, with noise and heat the undesirable byproducts, thus leaving only 1% of the energy available for size reduction. Coupled with this is evidence that many of today's crushing and grinding techniques not only are antiquated but have little or no scientific base of understanding." Throughout the past 50-plus years, grinding efficiency has been referred to as being from 1 % to 10%. The measurable basis for these observations is difficult to establish, but they appear to be based on production performance compared to theoretical energy requirements for the work done. The theoretical energy requirements were determined from the energy required to break a single particle or a thin layer of monosized particles by compression or by a single impact from an impact-testing device. King, Tavaras and Middlemiss (1997), in a paper titled "Establishing the energy efficiency of a ball mill" described such a test. These batch single-breakage tests generally do not duplicate the continuous-breakage process in the comminution machinery used in ore concentrators. Operating efficiency The currently available crushers and grinding mills are fed at a controlled continuous flow rate, and they impart the breakage pattern needed to produce the desired size reduc¬tion. While theoretically inefficient, they are the most practical comminution machines available. The operation of this machinery is relatively simple and can be controlled with available control systems. The essential factors in determining the operating efficiency of comminution machines are energy, as consumed in the machinery, and the required energy, as determined from test work. Bond work indices, obtained from bench-scale or pilot-plant crushability and grindability tests are used in the following Bond equation (Bond, 1960) to determine the energy required to produce the required size reduction

Jan 1, 1999

By Patrick M. James

Many in mining believe that the industry is slipping behind in public image, in public mineral education and, particularly, in mining's basic right to do business. It is critical to turn things around - to change the rules of the game soon - before mining loses any chance of recovery. In the industry's experience, there is a technique of project development and operation that can help gain ground on public opinion. Only recently has this technique had a name, and that is sustainable development. What is sustainable development? Sustainable development is broadly accepted and supported by the international business community. And significant thought and work has been devoted to the subject. Research into the topic was initially confusing because the definitions of sustainable development are written in terms the writer, and his or her constituency understands. To some, it refers to environmental protection. To others, it is about social responsibility. And to the developer, it is about economic benefit. Slowly a theme emerged from the research that seemed to be part of all the definitions. That theme has three parts. First, the development, whatever or wherever it is, must contribute economic value to the developer and the stakeholders of the development. Stakeholders in this context include employees, neighbors, residents of the surrounding area, citizens of the country, suppliers and all others affected by the development. Next, sustainable development must contribute environmental value. This means that at a very minimum it must be neutral and not cause irreparable damage. And last, and what is new for miners and most other businessmen, it must contribute real social value to the stakeholders. How to use the concept? With that three-part theme as a working definition, the mining industry must find ways to use that “the challenge is for an individual company to translate what sustainable development means for them. There is no one model..." Donald W. Gentry is president and chief executive officer of Polymet Mining and former professor emeritus of mining engineering at Colorado School of Mines. He added: "While simple to state, the concept is anything but simple to implement. Present trends indicate ... one wins the sustainable development argument by staking out the high ground early..." In other words, start using sustainable development principles at the beginning of the project. New ways of doing business The mining industry's challenge is to implement sus¬tainable development universally around the world. Paternalism as a way of doing business is dead and a new model of doing business is emerging. All types of businesses are learning to deal with the stakeholders of new developments. Mining cannot be an exception to these trends. To understand how the international mining business came to sustainable development, it is useful to look back a few years. For many North American mining engineers, there were few opportunities for foreign assignments from the 1960s to the 1980s. American companies had been asked to leave many of the Latin American countries. Africa was going through a continent-wide upheaval. The Soviet Union and China had Eastern Eu-

Jan 1, 2000

By Joseph M. Hans

INTRODUCTION The decontamination of a uranium mill site is principally an earth moving task. The task can be completed effectively and efficiently by selecting the proper earth moving equipment and using that equipment within its particular constraints for decontamination work. A considerable backlog of experience concerning earth moving equipment selection and use was acquired during the decontamination work conducted at the former Shiprock, New Mexico, uranium mill site. This experience was acquired by trial and error while decontaminating over a wide range of depths and different types of terrain. Many different types of excavation equipment were available to the Shiprock decontamination project because the Navajo Engineering and Construction Authority (NECA) heavy equipment operator's school was doing the decontamination work and a variety of equipment was needed for training. The equipment is described and its general use for decontamination is discussed. Its on-the-job use for decontaminating the Shiprock millsite is presented. [EARTH MOVING/ANCILLARY EQUIPMENT DESCRIPTION AND DECONTAMINATION CONSTRAINTS] Tractor-Scrapers The most commonly used scraper is the bottom-dump type because it is highly mobile and can dig, carry, and spread loads. In this type a tractor is connected with a yoke to a scraper bowl (Figure 1). The bowl is essentially a rigid box that pivots around the rear axle, opening to the front. A hardened steel knife, or bit, is attached to the leading edge of the bowl. When scraping, the bit penetrates and forces the earth overlying the bit edge into the scraper bowl as the scraper moves forward. When the bowl is full, it is closed by lowering a closure or apron against the cutting bit. The load is dumped while moving, by raising the apron and forcing the load from the bowl with a crowd, or ejector. The height of the dumped load is controlled by adjusting the bit height when dumping. Scrapers of this type generally do not have enough power to scrape up loads unassisted. Loading assistance is generally provided by a crawler tractor that pushes from the rear against the scraper's push block. Some scrapers are made self-loading by adding an additional engine for the extra power needed for the load pick up. Another commonly used self-loading scraper uses a loading elevator (Figure 2). The elevator consists of two roller chains carrying crossbars. The elevator is rotated when scraping and the crossbars cut the earth above or in front of the bowl bit, forcing the load into the scraper bowl. No apron is used on elevating scrapers. The load is dumped by partially retracting the bowl floor and pushing with an ejector. Scrapers have some undesirable characteristics when used for decontamination work. One characteristic is that a clearance space exists around the apron when it is in its closed position. Part of the load (especially sandy soils) dribbles through the

Jan 1, 1981

By R. F. McFarlin, H. E. El-Shall

Since its inception, the Institute of Phosphate Research has examined every promising proposal in order to find an economical, practical and environmentally sound method of dewatering phosphatic clays. The in-house development of the FIPR/DIPR process (DIPR = Dewatering Instantaneously with Pulp Recycle) would appear to answer that quest. When small amounts of fibrous materials are mixed with clay slurries and flocculent, rapid dewatering occurs. Clay solids contents of 20 to 25% can be attained by simply rolling the clay mass down an inclined screen. Higher solids (>40%) are reached by physically pressing the cake; e.g., with a belt press. Because mechanical pressing devices are too expensive for phosphatic clay treatment, FIPR has developed the "sandwich" method of simultaneously dewatering the phosphatic clays and reclaiming the mine cut, as well as a second alternate, the "median strip" proposal. Economic projections of these bench-scale alternates suggest that the FIPR/DIPR process should be no more costly than present practice, when credit is given for the simultaneous reclamation accomplished.

Jan 1, 1993

By W. Jacobi, K. Eisfeld

The primary objective of dosimetric modelling for the inhalation of radon daughters is the evaluation of the dose distribution and of the mean dose to sensitive cells or target tissues in the respiratory tract which are assumed to be responsible for the initiation of lung cancer. The results of such dosimetric models are influencing the practice of radiation protection in a twofold manner: Firstly, they enable us to identify appropriate operational quantities for air monitoring and for the surveillance of individual exposures to radon daughters. Secondly, they can be used to derive secondary and operational limits for these radionuclides from the basic dose limits. The dose to target tissues in the lung from inhaled radon daughters depends on various physical and biological parameters: The unattached fraction of daughter atoms, the AMAD of the carrier aerosol for attached atoms, the geometry of the respiratory airways, the breathing conditions, the parameters for the retention and translocation of daughter atoms in the different lung regions, and the location of the target cells or target tissues, respectively. Therefore, dosimetric modelling for radon daughters has to include a sensitivity analysis showing the influence of these parameters. To meet these objectives we have recently described in a comprehensive report an improved dosimetric model for the inhalation of 222Rn, 22ORn and their short lived daughters (Jacobi and Eisfeld, 1980). In the following this model is shortly described and the main results are summarized. DOSIMETRIC MODEL [Model Structure]. For the geometry of the bronchial airways the model A of Weibel (1963) was used. Each bronchial generation i (i =0.1 .... 16) was considered as a separate compartment i, for which the following transfer pathways for radon daughter atoms were taken into account (see fig. 1): (1) the uptake by deposition (deposition probability wi); (2) the translocation by ciliary clearance (rate constant [A]i); (3) the transfer to blood by diffusion through the bronchial epithelium (rate constant [A]b). To distinguish between attached (Index a) and free (index f) daughter atoms each bronchial generation was divided in two subcompartments (i.a and if) with a transfer rate constant [a]s (ia- if) characterizing the desorption or dissolution rate of attached atoms in the mucus fluid. Thus for daughter atoms primarily deposited in attached form the transfer to blood is a twostep process with the rate constants [a]s and [A]b which were assumed to be independent of the generation number i. [Deposition]. The deposition probabilities wi in each bronchial generation were calculated on the basis of the laminar diffusion model of Gormley et al. (1949); however the deposition in the upper airways was corrected for turbulent air flow, taking into account the results of our deposition experiments (Martin et al., 1972). Figure 2 shows the resulting deposition probabilities per unit surface area of the bronchial generations for nose breathing at a rate of 20 1/min. The values for attached daughter atoms refer to an AMAD of 0.2 - 0.3 µm ([O]g =3) for the carrier aerosol; the deposition values for free atoms were derived with a diffusion coefficient of 0.05 cm2/s, assuming a prefiltration of 50% in the naseopharyngeal region.

Jan 1, 1981

By J. F. Rakszawski, Robert P. D'Orazio

The use of cyanide in the processing of gold has been documented for over a century. It is still one of the more effective methods of isolating gold, as witnessed by the fact that in 1987 cyanidation was used in the processing of 91.5% of the gold produced in the United States.

Jan 1, 1990

By E. De Souza, H. Heidrich

In underground operations the evaluation of blasting fume dilution times and the application of recirculation techniques require accurate determination of the concentration of toxic fumes produced by the explosive used. This paper describes a methodology for obtaining concentrations of detonation products from commercial explosives tested under typical rock confinement and looks into the development of techniques for reducing toxic gases produced by explosives. This paper also analyses the application of techniques for the removal of toxic gas which would enable reduced re-entry times, a reduction in the quantity of fresh air and the use of air recirculation techniques, thereby providing substantial economic benefits to the industry. INTRODUCTION Toxic fumes from blasting operations are among the most common and serious hazards encountered underground. The implication of high concentrations of toxic gases emitted from blasting on the mine ventilation system are very significant and, in many cases, catastrophic results have occurred. Because blasting is a routine practice underground, the mining engineer is always confronted with the problem of controlling the fumes within acceptable levels. Typical control procedures employed underground include preventing or reducing the amount of fumes liberated from blasting through the proper selection of explosives and proper blasting techniques; removing the fumes through local exhaust systems or auxiliary ventilation (dilution ventilation); blasting off-shift, or at restricted times; localizing or isolating the effects of blasting where workings are isolated or sectionalized in the ventilation system; and absorbing some of the components of the fumes using air-water sprays. Dilution is one of the most effective control measures against blasting fumes and is universally applied. However, for precise determination of the dilution times for the fumes to reach safe concentrations, accurate determination of the concentration of toxic fumes produced by the explosives used is required. Research work has been carried out in determining the types and volumes of gases produced from blasting. In the case of ideal detonations, the nature and quantity of the fumes produced can be predicted by thermo-hydrodynamic calculations. However, commercial explosives exhibit non-ideal behaviour which is enhanced by the relatively small diameters of application in underground mines. The confinement offered by the surrounding rock further complicates the situation. It is well recognized that confinement affects performance and consequently fume generation. A detailed data base of fumes generated by commercial explosives at various conditions is introduced and the effect of blasting parameters such as, confinement, shock desensitization and incomplete detonations on toxic fume production is described for a variety of commercial explosives in use today. Thermodynamic predictions on fume production by the explosive compositions tested are also presented and compared with experimental results and their significance in assessing the quantity of fumes produced is discussed. A mathematical model capable of determining dilution ventilation requirements and dilution times to reduce the fume concentrations in the blasting zone to acceptable levels has been developed and it is being incorporated into a network system so that the effects of blasting can be simulated in typical mine ventilation systems. Solutions to the problem of controlling blast fumes within acceptable levels are described. These solutions include the use of control procedures which prevent and reduce the amount of fumes liberated from blasting. The use of low fume explosives, proper blasting techniques, dilution ventilation and filtration can be combined to effectively control fume production and save on costs. The application of these techniques to enable recirculation of underground air with reduced quantities of fresh air requirements is evaluated in a case study.

Jan 1, 1993