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By David S. Smith

Almost nine years after the assertion of enormous and pervasive federal control over coal mines by means of the Federal Coal Mine Safety and Health Act of 1969, 30 USC Section 801 et seq, as amended by the Federal Mine Safety and Health Act of 1977 (which basically increases the punitive powers of government) a brief assessment of some of the underlying causes, present status, and possible alternatives to governmental control is appropriate. In a governmental system based upon the principles of democracy and the protection of individual freedoms and constitutional rights the ultimate mechanism for abrogation of these rights is the creation and exploitation of fear. Machiavelli in rendering advice concerning the power over men provided by fear stated, "... it is much safer to be feared than loved .. fear preserves (the government) by a dread of punishment which never fails." The Prince. Within the context of the assumption of control by the government over industries or segments of our population - popular fear provides a convenient and almost necessary mechanism for the avoidance by law or regulation of democratic principles. The popular fear of coal mines has been successfully exploited by various goal oriented groups in the past and most recently by a Congressional constituency with a vested interest in controlling the mines through regulation.

Jan 1, 1979

By B. Karatzoglou

The penetration of sustainability principles in business practices has introduced the necessity for a simultaneous integrated assessment of the corporate economic, environmental and social performances. Continuous technological, political and legal evolutions indicate that Sustainable Development (SD) principles will soon have to be implemented in all sectors of the Greek corporate reality, among them in the Mineral and Mining Industry. However, the ultimate criteria for the evaluation of a company?s performance remain its profitability and its market value. Managers, shareholders, financial intermediaries, all alike, emphasize and base their decisions on the various accounting produced ratios of return (ROA, ROE, ROCE, Residual Value) included, or produced by data found, in the financial statements of the companies. This paper offers evidence that the traditional accounting ratios discourage the implementation of investment plans that aim at improving the environmental performance of companies and therefore can be inadequate and misguiding for SD applications. International studies have consistently traced a positive relationship between the environmental performances of firms and their financial positions. Yet, the fact that investors tend to reward proactive companies by increasing their corporate values is disregarded by the return ratios, which usually set up the basis for management rewards and bonuses. The book value of the firm assets and owners? equity from which such ratios are produced diverges strongly from the real corporate value, defined as the traded amount for the stock of a firm. There is an urgent need for an adjustment of the return ratios so that they can record the positive economic impact of a more sustainable attitude and encourage decision makers towards this direction.

Jan 1, 2003

By C. L. Pillar

In many cases, substantial time may elapse between the exploration stage and a serious detailed planning stage which has the objective of putting a mineral property into production. This is always true where properties change ownership and new partners are brought in for joint venture situations. Such conditions make it imperative that the recording of complete and accurate geological data for the property is done during the active part of the exploration phase and accomplished by competent geologists and engineers. This work should not be left to helpers and technicians. Furthermore, it should not have to be redone years after the actual drilling was accomplished, when either those involved have parted company or those present have lost the touch and feel of their part of the exploration program and must rely on a memory sketchy in detail, at the best. All of this emphasizes the importance of thorough geological evaluation and reporting during the exploration program. To evaluate and comment on the application of geological data by a mine planning department, we must assume (1) that sufficient diamond drill holes have outlined the existence of a potentially economic ore deposit, and (2) that the next step is to obtain the required information to allow a detailed feasibility study. This paper will comment on: 1. what types of geological information are required, 2. how it can be obtained, and 3. comments on the application of this information by the mine planners.

Jan 1, 1973

By R. Q. Phillips

INTRODUCTION Since Canada commenced sulphur exports in the early 1960's, the volume of exports has grown steadily and by 1984 Canadian sulphur accounted for 45% of internationally traded sulphur. In 1984 a total of 7.6 million tonnes of Canadian sulphur was shipped to offshore markets and the U.S.A. Total shipments, including domestic sales, were 8.3 million tonnes. With a production of some 5.6 million tonnes, inventory was drawn down by nearly 3 million tonnes in order to keep pace with demand. Since 1979, world consumption has exceeded total world production and Canadian sulphur inventory has played a truly crucial role in balancing world supply and demand. Consequently, by the end of 1984 the Canadian inventory, which stood at 21 million tonnes in 1978, had been almost halved to 12 million tonnes. This inventory, however, is still by far the largest in the world and will continue to play a vital role in the world supply until expected depletion around 1990. Canada is the world's second largest producer of sulphur following the U.S.A.. The Current production level of between 5 and 6 million tonnes per annum is forecast to be maintained to 1990 with only a marginal fall in that production level in the following decade to the year 2000. This will ensure that Canada retains its position as one of the largest producers and most important suppliers of sulphur in the world market well into the future. CANADA AS WORLD SUPPLIER OF SULPHUR Since Canada commenced sulphur exports in the early 19601s, exports have grown steadily. (Figure 1) The rate of growth accelerated in recent years and in 1984 a total of 7.6 million tonnes was sold for export, compared to 3.0 million tonnes in 1975 and less than 1 million tonnes in 1964. World trade in sulphur also increased during this period, from 11 million tonnes in 1975 to 17.0 million tonnes in 1984 (Figure 2). Most of this increase in volume was accounted for by Canada and Poland. As a result, the importance of Canada as world supplier expanded from 28% in 1975 to 45% in 1984. The markets for Canadian sulphur can be divided into two broad categories, i.e., solid and liquid (Figure 3). The solid market is the largest, taking about 70% of Canadian output, virtually all of which is sold offshore. The U.S.A. and the Canadian domestic market take about 20% and 10% respectively, in liquid form. In 1984, (figure 4) total Canadian sulphur sales exceeded 8 million tonnes with offshore export sales at a record 5.8 million tonnes and sales to the U.S. of 1.8 million tonnes. Domestic sales accounted for only 0.7 million tonnes. The general trend for offshore export sales has been upward whereas there has been relative stability in domestic sales. The exceptional growth in offshore sales in the 1970's necessitated considerable capital investments by the industry in order to establish a logistics system capable of handling increasing offshore tonnages. Over 50 sulphur plants are currently operating in Alberta and British Columbia. Some of these are connected by rail, (Figure 5) the balance being generally connected by road. There are currently fourteen of the larger rail-connected plants designated as export plants, and these are equipped with sulphur forming facilities. The long distances between these plants and tidewater terminals in Vancouver is a very significant factor in complex logistics of Canadian offshore sulphur exports (Figure 6). Alberta and B.C. plants are located an average of 1,400 kilometers from tidewater.

Jan 1, 1986

By Nasib Al Habib, Hooman Askari-Nasab, EUGENE BEN-AWUAH

In-pit crushing and conveying (IPCC) is getting popular in deeper open pit mines as a suitable alternative to truck haulage because it offers a lower operating cost due to shorter haulage distance and less truck requirement. Semi-mobile in-pit crusher, currently the most popular IPCC system, is relocated every two to five years, and the short-term plan needs to be updated accordingly. To the best of our knowledge, short-term planning with IPCC is an area of research that has yet to be explored extensively, and hardly any model can generate a shortterm extraction sequence considering an IPCC. This research work proposes a mixed integer programming model to develop a short-term production plan within a time horizon of 12 months. The model's objective is to optimally allocate shovels to minimize the cost of material handling and maximize revenue subject to meeting plant requirements, maximum allowable tonnage variation and IPCC location constraints to achieve production and NPV targets set by the strategic plan. The proposed model will be implemented in an iron ore mine case study for verification. The model is developed and solved using MATLAB. The comparison of results between scenarios with and without IPCC justifies the use of IPCC in large open pit mines from a short-term perspective.

Jun 25, 2023

By Theodore H. Eyde

The porphyry copper province of the southwestern United States and northern Mexico, is the largest copper producing region in the world. Most of the deposits occur in Arizona which now produces over 54% of the newly mined copper in the United" States. As the many recent discoveries are brought into production, Arizona's share of the nation's copper production will increase. The term "porphyry copper" as used in this paper includes all deposits of generally low grade disseminated copper mineralization amenable to bulk mining techniques. Therefore, copper deposits such as Pima-Mission, Twin Buttes, and Vekol where the copper mineralization occurs principally in the intruded rocks will be considered as "porphyry copper" deposits. Most of the deposits have been discovered since World War II, consequently, the southwest, and particularly southern Arizona is perhaps the most intensively explored area for copper deposits on earth. Most of the major natural resource companies in the United States and Canada maintain exploration offices in southern Arizona. To illustrate the success of exploration efforts, the location of porphyry copper deposits either known or producing prior to World War II are shown on figure 1 and the deposits discovered since World War II which are either in production or in the exploration and development stage are shown on figure 2.

Jan 1, 1973

By Brian E. Davies, Bobby G. Wixson

INTRODUCTION The mining, extraction and refining of metals into useful materials is an honorable profession with an ancient and productive history. Civilizations have long depended upon metals for their implements, structures and machinery necessary for survival, growth, and development. Today, as in the future, metals and minerals are essential to the progress of mankind as we approach the 21st century with technologies and concepts that will require new concepts and production from the metal mining and processing industries. Today is the tomorrow we talked about yesterday and the implementation of new ideas is essential if we are to achieve "Frontier Technology in Mineral Processing". Unfortunately, the state-of-the-art mining, processing, smelting and refining methods historically employed were less than desirable environmentally, which has caused todays public to view the metal mining and processing industry with concern. Presently the ancient rock and chat tips connected with old Roman lead-zinc mines which were worked in Wales are causing stream contamination problems that must be corrected at considerable expense to the local governmental agencies (Firth, J.N.M., et al., 1981). However, improved progress has been made and the environmental control and clean-up procedures utilized by modern lead-zinc operations in the "Viburnum Trend" or "New Lead Belt" of South East Missouri may serve as a model to illustrate how alternative technologies, environmental concerns and cooperative efforts contribute to efficient mineral production within the confines of a U.S. National Forest region without significant environmental impacts. This area received international attention when it was opened within the boundaries of the Clark National Forest in the mid 1960's. By 1970 it ranked first in the world by producing 432,576 short tons of lead ore and has since continued in this position producing 451,153 short tons of lead ore in 1983 representing 91% of the total United States lead production (U.S. Bureau of idines, 1983). This production comes from a mining area approximately 40 miles long by 10 miles wide extending in a north-south direction. Geologically the lead is disseminated throughout the Cambrain age Bonneterre Formation, mostly a dolomite, at depths ranging from 700 to 1200 ft. (213.4 to 365.8 m). Galena (lead) is the principal ore mined with lesser quantities of sphalerite (zinc), chaleopyrite (copper) and silver being recovered as economic co-products. This new mining development in one of Missouri is most scenic areas could have led to a classic confrontation between industries and environmentalists. Instead, through the cooperation of industries, governmental agencies and university researchers, the Ozark environment has been protected, and may serve as a model for future coexistence of the mineral industry with the environment. Cooperative Program Development As early as 1965 five new mines and two lead smelters were being prepared for production and environmental concern was an essential part of that planning. All the mining companies became acquainted with representatives of the various federal and state agencies and scientists from the University of Missouri-Rolla (UMR). During the next six years, St. Joe Minerals Corporation, American Metal Climax, Inc., Cominco American Incorporated, Ozark Lead Company (Kennecott) and American Smelting and Refining Company participated in cooperative research efforts. The companies also participated in information exchange meetings, the development of improved analytical methods and the utilization of research findings in the modification or construction of pollution-abatement facilities. Early on research was carried out to determine background values and to establish natural baselines of trace metals in the surrounding environment. The unusual topography and drainage pattern of the mining area were helpful since it was possible to establish control sites and test sites below individual mine-mill operations to study waste-water effects on the receiving streams. The location of the ‘New Lead Belt', mines, mills,

Jan 1, 1985

By Tim O’Neil

Until about three years ago, Asarco's copper business consisted predomi¬nantly of custom smelting of ores and concentrates produced by other mining companies. Since then, the company has been transforming itself into a fully integrated producer in copper mining, smelting, and refining. In doing this, Asarco has lowered its costs and restructured its operations and finances. Now, Asarco hopes to complete this process by spending $260 million over the next three years, to expand and modernize its copper facilities and boost production by some 40%. The capital spending program includes: • $130 million at the Ray mine about 113 km (70 miles) north of Tucson, AZ, to expand mining capacity and install an in-pit ore crusher, mill, and concentrator; • $100 million, to expand the mining capacity at the Mission Complex located south of Tucson, AZ, and refurbish the adjacent, idle Pima mill and concentrator. (Asarco had earlier exercised an option to purchase the Pima mill for about $6 million); and • $30 million, to modernize the copper smelter in El Paso, TX, with a new flash smelting process. The new program is in addition to a recently completed first-phase expansion of the Mission mine, mill, and concentrator. This $13 million debottlenecking of existing operations in¬creased production capacity by 46% or 24.5 kt/a (27,000 stpy), to 79 kt/a (87,000 stpy) of copper in concentrates. In addition, a previously an¬nounced $12 million expansion program at the Ray mine is scheduled for completion in early 1990. It will expand mill capacity, to offset the anticipated effects of increasing ore hardness as the pit deepens. Ray produces 68 kt/a (75,000 stpy) of copper in concen¬trates and an additional 36 kt/a (40,000 stpy) of electrowon copper. By 1992, when Asarco's expanded and modernized copper facilities are operating at capacity, the company's mine output will have increased by 67 kt/a (74,000 stpy), to 263 kt/a (290,000 stpy) of contained copper. That will be enough to provide all of the feed required for Asarco's two copper smelters - by then, both of them modern, state-of-the-art facilities. Asarco's expansion and modernization program will further reduce costs and provide added assurance that the company's copper business will be profitable at the bottom of the cycle, according to Chairman Richard de J. Osborne The Ray mine portion of the new program will include construction at the mine of a new mill and concentrator, with a capacity of 18 kt/d (20,000 stpd) of ore. These new semiautogenous grinding mills and large capacity flotation cells will augment the present 27.2 kt/d (30,000 stpd) concentrator located 29 km (18 miles) away, in Hayden, AZ. A 29 km (18 mile) pipeline will be built to carry tailings in slurry form from the new Ray mill to the present tailings pond. Concentrates from the new facility will be shipped by rail to Asarco's Hayden copper smelter. Ray's present 27.2 kt/d (30,000 stpd) crusher, adjacent to the open-pit, will be replaced by a 54.4 kt/d (60,000 stpd), portable in-pit crusher and conveying system. This will reduce the more expensive ore haulage by truck. The Ray project is scheduled for completion by 1992. It will increase the mine's annual output of copper in concentrates by an additional 33.5 kt/a (37,000 st), to 102 kt (112,000 st). And the project could mean an additional 400 jobs at the 730 employee Ray operation. Work at the Mission Complex involves reactivation of the Pima mill and concentrator and expansion of mine output sufficient to provide ore to both the Pima mill and the present Mission mill. In 1985, Asarco purchased the Pima mine, which occupies one end of the Mission pit. The work at Mission will increase its annual capacity by 33.5 kt (37,000 st), to 112 kt/a (124,000 st) of contained

Jan 1, 1989

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Shrinkage stope mining can be employed in steeply dipping veins where both the wall and the vein material are sufficiently strong to stand with only minimal ground support. Shrinkage stope mining takes advantage of the swell in volume produced in breaking in-place ore. This expanded volume of ore is allowed to occupy part of the mined-out vein, and in so doing provides ground sup¬port and a working surface for the miners. Only enough of the ore is drawn after each stoping cut to allow work¬ing room. Not until the entire mining block has been broken is the full amount of broken ore drawn from the stope. DEVELOPMENT PRACTICES FOR SHRINKAGE STOPING The vein is developed into stoping blocks, each of which is 76 to 91 m (250 to 300 ft) in length and 61 m (200 ft) between levels. Level development, described later, is either on the vein or parallel to it with crosscuts periodically driven over to the vein. There are usually two exits from each stope. The first entry is a raise at one end driven either conventionally by raise climber, or by a raised borehole. Present practice is that this raise, put in from level to level, is either a 2.1- x 2.1-m (7- x 7-ft) Alimak raise or a 1.2- to 1.8-m (4- to 6-ft) diam borehole. Either raise system will require pro¬visions for an ore chute carried with the stope to the haulage level or the scram drift. At 9- to 12-m (30- to 40-ft) vertical intervals, dogholes are developed from the raise so that when stoping breaks back into these dogholes, timber or steel sets and necessary service facilities can be installed. The second entry to the stope may be either a similar raise at the stope endline, or more economically, a cribbed raise carried with the stope. As previously stated, development of the ore block is achieved by either drifting on the vein or by paralleling the vein with a footwall lateral and then crosscutting over to the vein at specified intervals. In the following discussion, three basic cases are illustrated in Figs. 1, 2, and 3. In Fig. 1, development is in the footwall by a 3- x 3-m (10- x 10-ft) trackless lateral. A heading is also driven on the ore and taken either full width or driven a standard size and slabbed to the vein width. Approxi¬mately 9-m (30-ft) long crosscuts on 7.6-m (25-ft) cen¬ters are driven from the lateral to the vein. For ease in muck handling they are usually put in on a slant rather than at right angles. A single orepass, located off the lateral and in the center of one stope block, can serve not only that stope, but also one on either side, up to a tramming distance of not more than 137 m (450 ft). Raises are put at the stope endlines on 91-m (300-ft) centers. The back of the vein is shot down and just enough broken material is drawn off to permit access for men to work off the broken ore. Development in Fig. 2 progresses by driving a 3- x 3-m (10- x 10-ft) trackless heading on the footwall side of the vein. The ore block is delineated by raises on 91-m (300-ft) centers. Crosscuts are again driven on 7.6-m (25-ft) centers over to the vein. Approximately four rounds are then taken up to the vein and each round is successively belled out so that all the crosscuts (pockets, in this case) connect. In this development, as opposed to that in Fig. 1 where there were no belled

Jan 1, 1982

By E. Steel, V. S. Znamensky, A. Wylie

The asbestiform habit is most commonly developed in certain amphiboles and chrysotile, but other minerals also may crystallize with this unusual habit. The habit may be characterized by (1) a fibril structure, single or twinned crystals of very small widths (generally less than 0.5 pm), which have grown with a common fiber axis direction, but are disoriented in the other crystallographic directions; (2) anomalous optical properties, primarily parallel extinction; (3) unusual tensile strength; (4) high aspect ratio; and (51 flexibility. In addition, there is evidence to indicate that some amphibole asbestos may have unusual surface properties. ASBESTOS AND THE ASBESTIFORM HABIT Asbestos is defined as a group of highly fibrous silicate minerals that readily separate into long, thin, strong fibers of sufficient flexibility to be woven, are heat resistant and chemically inert, and possess a high electric insulation, and therefore are suitable for uses where in- combustible, nonconducting, or chemically resistant material is required (Gary, et al., 1974). The most common minerals that may occur with the asbestiform habit are chrysotile, grunerite (amosite], riebeckite (crocidolite), actinolite, anthophyllite, and tremolite; although the development of this habit among these minerals is rare. Other minerals, most notably other amphiboles, can occur in this habit, but no others have been mined commercially as asbestos. All asbestos is confined to metamorphic rocks, even though other habits of the amphiboles are common in igneous rocks. Slip fiber veins are the most common deposits (South African amosite, Canadian chrysotile, etc.), but mass fiber deposits also may occur such as California chrysotile or mountain leather. Field relations and experimental data support the hypothesis that metasomatism is the dominant process in the formation of asbestos fibers, amphibole as well as serpentine. The crystal habit of a mineral is the shape or form a crystal or aggregate of crystals take on during crystallization. The asbestiform habit has a number of characteristics that differentiate it from other habits. Chief among these is the fibril structure.Afibri1 is a single or twinned crystal with a very small width, generally less than 0.5 pm, and an extremely high aspect ratio; bundles of fibrils may have lengths reaching into the cm. Fibrils share a common crystal growth direction along the long axis of the fiber, but appear to be disoriented with respect to one another in the other crystallographic directions. The structure of the individual fibrils, and the organization of fibrils within a fiber may differ among the various types of asbestos. The fibril structure of asbestos is probably a factor in controlling a number of secondary properties that include high tensile strength, flexibility sufficient for weaving, and anomalous optical properties. The high tensile strength of asbestos is quite remarkable, exceeding that of the ordinary varieties by a thousand fold (Zoltai, 1978). In part, this may be attributed to the lack of defect on the fibril surfaces (Zoltai, 19781, but some contribution also must come from the existance of bundles and the nature of the ordering of and forces among the fibrils. The flexibility is not only enhanced by the fact that the fibrils may slip past one another, but by their small widths and extreme aspect ratios. However, structural defects parallel to the fiber axis also may contribute to this property. Among the common commercial asbestos types, the greatest flexibility is found in chrysotile and the least in grunerite asbestos (amosite), varying inversely with fibril thick- ness. In addition, at least for chrysotile, composition also may affect the flexibility of the fibers. Finally, most asbestos minerals are mono- clinic, but optically display parallel extinction (Heinrich, 1965, and Wylie, 1978). This also must be due, in part, to the ordering or lack of ordering of the fibrils. The fibril structure is probably a random

Jan 1, 1986

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 David A. Lindsey

In this paper, we present a physiographic classification of gravel deposits, identify some field measures of aggregate quality, and describe sampling strategies for mapping aggregate quality in the Front Range Urban Corridor of Colorado. Gravel occurs in four types of deposits along the Front Range of Colorado: (1) piedmont alluvial fans, (2) high dissected terraces, (3) high (continuous) terraces, and (4) floodplains and low terraces. The four deposit types are the basis for definition of gravel-deposit models, which differ in geologic environment, deposit characteristics, deposit size, environmental impacts of mining, and post-mining use. Of the four types, floodplains and low-terrace deposits have been exploited most for gravel. High terraces also have been mined. Deposits of high dissected terraces and dissected alluvial fans have produced minor amounts of gravel, mostly for use as local fill. Studies to identify useful field measures of aggregate quality and to determine the appropriate scale for sampling were conducted on gravel deposits beneath the flood-plain and low terraces of the South Platte River north of Denver and the Cache la Poudre River between Ft. Collins and Greeley. Gravel deposits of the South Platte were selected for study because they represent the last major resource of gravel for aggregate in the Denver metropolitan area. The Cache la Poudre valley is the primary source of gravel aggregate for rapidly urbanizing areas around Ft. Collins and Greeley. Of all the field measures of gravel quality, the most important is particle size. Various parameters that measure particle size were investigated, but the easiest to determine, and the most useful to the aggregate industry, are measures of the proportion of coarse sizes. Particle size varies widely in gravels of the urban corridor. Gravel lithology, particularly the presence of deleterious particles such as reactive volcanic rocks and decomposed rocks, must be determined to assess aggregate quality. Studies of South Platte and Cache la Poudre gravels have not revealed large proportions of deleterious lithologies. Studies of particle shape revealed only small numbers of physically weak shapes in both the South Platte and Cache la Poudre gravels. Particle shape does not appear to vary from place to place in the gravels. Rounding, or the ab¬sence of angular corners, may play a role in the tendency of loose gravel to yield under load. South Platte and Cache la Poudre gravels are rounded, and rounding tends to increase upstream. Before the quality of gravel in a valley can be assessed, the stratigraphy of the gravel must be known because it affects sample design. If only a single gravel layer is to be studied, only the thickness of the gravel must be known before sampling. If, however, more than one layer of gravel-distinguished by particle size, sorting, or lithology-is present, sampling must take into account the thickness and distribution of each layer. For example, three gravel layers of different particle size underlie the floodplain and low terraces of the South Platte. The tributary Cache la Poudre, on the other hand, contains gravel layers that can be distinguished only locally and vary slightly in particle size. Sampling experiments were conducted on gravel bars in the South Platte River and in gravel pits on the floodplain and low terraces of the South Platte and Cache la Poudre Rivers. Experimental results indicate that, even where sample error is large, regional and stratigraphic variation may be measured if it is large relative to sample error. Using vertical trench samples weighing an average of 40-90 pounds, down-stream decrease in coarse particle size was mapped for Cache la Poudre gravel, and differences among superposed gravel beds were measured for South Platte gravel.

Jan 1, 1999

Disposal of solid and liquid wastes from mining and milling operations has become a focus of attention in recent years. Consequently, the geotechnical engineer is more frequently associated with design of disposal facilities. Jack Wulff, a representative of one consulting firm with experience in this field, talks with ME about the geotechnical engineer's contribution. Mr. Wulff, mine and mill waste disposal today is receiving more attention-and many more dollars-than in the past. Regulatory agencies have promulgated demanding requirements concerning mill waste (especially uranium mill waste) and specialist consultants are frequently being called in to design disposal systems. What can the experienced geotechnical engineer add to the development of waste disposal facilities? It's true that geotechnical engineers have recently become more visible in the minerals sector, but we've been around mining for a long time-as foundation engineers for mills and mine support buildings, and for geotechnical aspects of open pits, pipelines, railroads, roads, and so forth. We've also designed water supply dams, which could be considered recursors of tailings dams and other forms of disposal systems. Let me show by example what geotechnical engineering can contribute to waste disposal facility design. I recently was asked to look at several small wastewater impoundments which had failed. The design of the earth embankments appeared, at first glance, to be rather conservative. They were only about 3.4 m high, the slopes were 3 to 1 (horizontal to vertical), with about 24 m of bottom contact with the earth foundation. But there had been three episodes of dike failure in a few months, all upon initial filling. The problem was the nature of some material in the foundation that was also used as structural fill-a dispersive clay that readily deflocculated and eroded on contact with water, especially with seeping or flowing water. The design did not include the defensive measures required to use such materials confidently in water retention structures. Apparently, the designer did not recognize the potential problem posed by the unusual soil characteristics. That's the sort of engineering information fundamental to geotechnical engineering, but not necessarily a part of the civil, structural, or mining engineer's professional training or experience. This difference in expertise is not simply a matter of factual knowledge-it also affects the design process. The structural engineer can design a structure to fit the client's functional needs and can specify the materials (grade of steel, PVC, concrete, etc.) to meet the demands of the design. The geotechnical engineer's design process is somewhat different, since we work with natural materials, not man-made materials produced according to specification. We first have to establish what kinds of materials are available (usually directly from the chosen site, since importation of materials becomes expensive). Then we determine the engineering properties and characteristics and, finally, design the structure to fulfill the client's requirements within parameters dictated by available materials. In brief, the geotechnical engineer works with naturally-evolved, not man-made, materials and uses design techniques developed to accommodate that fact. Because no two sites are exactly alike in configuration, structural (geologic) characteristics, or available materials, there will be unique design and judgmental questions to be resolved on virtually every project. You said that water retention dams could be considered the precursors of tailings dams-you probably had a specific technical relationship in mind. There's a very strong carryover of methodology from earth dam design, where the technology developed and where it has become very refined. In designing a tailings impoundment or similar facility, essentially the same procedures of field sampling are involved (with a geotechnical engineer or an engineering geologist directing the program since representative

Jan 7, 1980

By Leonard Harris

Good morning ladies and gentlemen. I would like to thank Hans von Michaelis for inviting me here today to talk to you on the subject of "How to Reduce Capital and Operating Costs at Heap Leaching Projects". The comments I have to offer are based on my experiences in gold heap leaching by Newmont Gold Company in Nevada, by Newcrest in Western Australia (formerly Newmont Australia), by Minera Yanacocha in Peru, by Penmont in Mexico, by Z­Mont in Uzbekistan, by the Hochschild Group in Peru and expe­riences in copper heap leaching by Magma Copper Company in Arizona and as a consultant on several other smaller projects worldwide. These comments are a mixture of technology, ideology, philos­ophy, strategy, logic and opinions and might be looked at by many of you as "one man's view". I like to look at them as "a view from the bridge". Let me add that it doesn't bother me that you may not agree with my comments and that I would be interested in your comments or criticisms at the end of this presentation. I don't intend to discuss the details of the mineralogical and metallurgical test work. I will assume, for this talk, that the test work has been conducted very carefully and that the ore in ques­tion has been determined to be heap or dump leachable. Let's begin with a couple of don'ts! ! First, "don't build the mill until you have a mine". This might sound to many of you as fundamental but believe me it continues to happen. The minable reserves should be well known and established as being economical to mine and leach. NPVs, ROis and IRRs are fine if you can be sure of the gold price you use in your calculations. The best method is to carefully estimate the capital cost and, most importantly, the operating cost per ounce of gold recovered which should include the repayment of the borrowed capital and the interest thereon. For example, if the operating cost under these conditions is $200 per ounce, you have a project. If it is $400 per ounce, walk away f om it. Incidentally, the same prin­ciple can be applied to projects involving other metals such as silver, copper, lead, zinc, etc. The second don't is - "don't be sheep". Don't climb on the bandwagon with respect to design and operation because every­one else does it a certain way. Keep an open mind, be versatile, look for change, look for every possible innovation for reducing capital and operating costs. Beware of those who will tell you "we tried that it won't work" or "it's too expensive" or "that's a crazy idea" , <' ''what do you know about mining - you are a metallurgist aren't you" or "what do you know about metallurgy - you are a miner aren't you" or "only geologists can handle this issue". Be careful about pigeon-holing people according to their basic degree and experience. In these days of low metal prices, we need to consider everybody's ideas, as crazy as some of them might seem.

Jan 1, 1998

By H. Nematollahi

The comminution laws The grindability of the ores or minerals is an important factor that could not be deduced from their mineralogical characteristics (Testut, 1958). Many theories have been proposed to determine the amount of work required in a comminution operation as a function of the nature of material, size and shape of the minerals before and after of comminution. The most important laws are proposed by Rittinger, Kick and Bond. According to Rittinger (1867), the work is directly proportional to the new surfaces produced during the comminution. Kick (1885) believes that the work needed to produce similar modification is proportional to the volume or mass of the material. Bond, during his 25 years of experience and research, reached to the conclusion that on the average the energy required for a known comminution is between the Rittinger&apos;s and Kick&apos;s laws (Bond, 1952). According to Bond&apos;s Third Theory, the work input is proportional to the new crack tip length produced during the particle breakage, and equals the work represented by the product minus that represented by the feed. The work input is given by the following equation: [111W = 1 Owi( FP FF(1)] where W is the The energy input to the mill, (kWh/st); F is the Feed size based on 80% passed, (µm); P is the Product size based on 80% passed, (µm); and W; is the Work Index, (kWh/st) Methods to determine the Bond Work Index Various methods have been proposed by Bond to determine the Work Index. There are some problems with the Standard Bond Grindability Test: •Requiring a special mill. •Requiring a relatively large amount of sample. •Being relatively time consuming and tedious. [ ] During the recent years, numerous attempts have been made to install faster and easier methods to determine the Work Index. A complete description of these methods is presented by Yap et al. (1982). Determining the Bond Work Index by a new size ball mill During the recent years, researchers at the University of Tehran determined the Bond Work Index, using the Anaconda simplified methods by its Denver 305- x 127-mm (12 x 5-in.) laboratory wet ball mill. The equation used to calculate the Work Index for the above mill containing 16.12 kg (35.5 lbs) of steel balls is as follow: [1 _ 1-1W;=0.54x [P [)(2)] In addition, an attempt has been made to determine the Bond Work Index similar to standard Bond ball mill procedure but by a different ball mill, smaller and in consequent portable and less laborious than the Bond ball mill. Method description Just as the standard Bond grindability test, this method is a locked-cycle dry grinding and screening process that is carried out until steady state condition is reached. Test equipment The test is conducted in a special ball mill, called the New Size Ball Mill (NSBM). The mill is scaled down with a coefficient of two-thirds from the standard Bond ball mill. It measures 200 x 200 mm (8- x 8-in.) inside (instead of 300 x 300 mm or 12 x 12 in. in Bond ball mill). Just like the Bond

Jan 1, 1995

By L. C. Raymond

Good afternoon. I have entitled my talk "Investment for Progress," implying thereby a spark of friendly competition by private enterprise with such government-inspired foreign aid and investment programs as the Alliance for Progress. While the subject of foreign investment is hardly a new one, the perpetual flux of the political and economic surface of this earth we inhabit makes it forever timely and forever interesting. Especially do tremendous opportunities for international investment constantly present themselves to the mineral industry, and just as often must the industry approach these opportunities with a great deal of investigation and caution. This is not to say that the mineral industry--no less than any other industry--should be unenthusiastic about the foreign scene, but like everyone who proposes to put his money where his hope is, it should consider all the factors and weigh the possible disappointments as well as the rewards.

Jan 1, 1964

By R. Cooper, K. De Wet, A. Janwong

"Horsehead Metal Products is a zinc refinery in Mooresboro NC treating crude Waelz Oxide (WOX) to produce Special High Grade Zinc by a hydrometallurgical process, comprising leaching, zinc solvent extraction/electrowinning (SX/EW), melting and casting and bleed treatment. Cadmium removal by zinc dust cementation plays an essential role in the bleed treatment circuit to prevent a build-up of cadmium in the circuit. Silica, although not present at significant levels in the WOX feed material (approximately 0.3%), had a significant impact on the efficiency of cadmium cementation with zinc dust, most likely due to the passivation of zinc dust particles with silica gel. Solid/liquid separation efficiency by filter press was also impacted. The silica levels were mapped throughout the circuit in an attempt to identify the source of the silica issues. Adjustments including using PLS as feed to cementation, pH control in WOX leaching and neutralization have been introduced to mitigate the issues created by silica gel in the circuit. INTRODUCTION Located in North Carolina, Horsehead Metal Products Inc. is a hydrometallurgical zinc refinery processing Waelz Oxide (WOX) to produce Special High Grade (SHG) zinc. The flowsheet, known as the Modified Zincex® process, consists of leaching, solvent extraction, electrowinning, melting and casting and bleed treatment as illustrated in Figure 1. Cadmium cementation, part of the bleed treatment process, provides an essential bleed of cadmium from the circuit, as metallic cement. The Cadmium cementation circuit consists of agitated cementation reactors, a zinc dust addition system, filter presses, agitated purification (releach) reactors and a cadmium concentrate filter press Silica can be found in several forms in acid leaching. The first form is fine silicate particles. The size of the particles can vary from one to twenty microns (1). Fine silicate particles can lead to the formation of crud in solvent extraction circuits (2, 3). At Mooresboro, crud has been observed in Depletion, which is part of the solvent extraction circuit. Silica content in the solution in this section is high. The second form of silica is monosilicic acid Si(OH)4, which begins to polymerize almost immediately upon being leached into solution. Monosilicic acid polymerizes to silica gel or a silica precipitate dependent primarily on supersaturation (5). Factors that influence the rate of polymerization include temperature, sulfuric acid concentration / pH, hydrofluoric acid concentration, seeding, and dissolved salts concentration such as zinc and aluminum in the solution (4, 5, 6). In sulfate systems, silica solubility is reported in the range of 150 to 450 ppm of SiO2 (3, 5, 7)."

Jan 1, 2016

By Donald M. Keagy

How many times have you said or heard someone say, &apos;What are you doing right now?" And how many times have you heard the reply, "Breathing air!" Probably never, or at most only a few times in a facetious vein. Breathing air is absolutely necessary to sustain human life, and yet most of us take it completely for granted. Beginning ten to fifteen years ago there was growing concern on the part of the general public that pollution of the air is a threat to man&apos;s health and welfare. Concurrently much thought and effort was devoted to the task of defining and resolving the problem of air pollution. In the course of this effort it has become quite clear that our National policy in this regard should be founded upon three basic axioms: (1) Action should not be postponed until emergency conditions make it compulsory; (2) The air cannot be purified after it is polluted, and (3) Air pollution control cannot be accomplished without close cooperation among all levels of government, industry, and the public.

Jan 1, 1967

By Syd S. Peng

Introduction Modern longwall mining, introduced to the US coal industry in the mid-1960s, is the latest coal mining technique. Today, longwall mining produces more than 15% of all underground coal production. The growth of longwall mining in the US is slow. However, nearly two decades of longwall mining have demonstrated that its benefits outweigh the disadvantages. Longwall Advantages Production and Economics - A typical longwall production ranges mostly from 0.9 to 1.8 kt (1000 to 2000 st) of clean coal per shift. This is about three to six times the production of comparable continuous mining units. Furthermore, there are much less rejects in longwall mining, typically 15% to 25%. It must be noted, however, that a new longwall with an inexperienced crew may produce less, depending on mining conditions. But as experience gains, production increases rapidly. Due to its high production, the cost per ton of coal mined over the mine&apos;s lifetime is cheaper with longwall mining than with continuous mining. A study on a conceptual mine of 25 year life using actual operational data (Dangerfield, 1981) concluded that the cost per ton of coal mined is 32% cheaper by longwall than by continuous mining. Another benefit from its high production potential is that it enables the production to be concentrated in a few longwall faces versus many units of continuous mining required to achieve a similar production. As a result, mine organization and management simplify considerably. Safety and Savings - Most US longwalls use shield supports that cover the roof with solid canopy and isolate a gob completely from the face area. Thus, if the supports are properly selected and operated, the potential for roof fall accident is almost zero. Furthermore, it is not unusual that a well-run longwall face is much more orderly and cleaner that most well-run machine shops. The ventilation system for longwall mining is the most ideal type. It is simple and unique. No auxiliary fans are required. The fresh air sweeps through the whole face. Its volume and velocity can be adjusted as demanded. Recovery - In spite of multiple entries layout for US longwall panels, the longwall method recovers up to 40% more coal than the continuous mining method (Dangerfield, 1981). The typical recovery for longwall mining is 70% to 85% versus 45% to 80% for continuous mining. Versatility - Longwall mining has been employed successfully in seam heights ranging from 0.8 to 3.7 m (2.7 to 12 ft); in seam inclinations from horizontal to 35°; in overburden depths from 46 to 914 m (150 to 3000 ft); in single or multiple seam mining; and in uniform or irregular seam characteristics. For seam height less than 1.2 m (4 ft), the plow is used instead of a shearer. However, in spite of recent developments, the plow still is unsuitable for hard coal seams. Slow Acceptance How come longwall mining, with so many advantages, has not spread more rapidly? Undoubtedly, the major reason is the large capital investment required. Under normal market conditions, the capital required for a longwall panel ranges from $7 to $11 million. Another factor is the uncertainty of its applicability to any specific coal seam of interest where longwall has not been used or operated successfully. Although longwall mining elsewhere has proven its applicability to wide-ranging seam conditions, there have been several failures in the past two decades. Analyses of those cases show that most failures can be attributed to inexperience, both operational and technical, including equipment selection. Therefore, personnel training on the longwall mining technique is absolutely necessary before its initial operation. Some earlier statements about adverse surface subsidence associated with longwall mining were misleading. Research has proven that surface subsidence under longwall mining is controllable and predictable. And, if longwall mining was mandatory since the onset of coal mining in the US, there would be no surface subsidence problems over abandoned mined land today. It is a blessing that surface subsidence occurs

Jan 3, 1985

By D. P. Bryan

The identification and-develop¬ment of new construction aggregate resources, the most common mineral mined, is an essential need in this country&apos;s growing urban areas. In developing an aggregate exploration program, a thorough geological identification of all potential aggregate resources in an area is still of primary importance but other factors have also to be considered. These include physical properties, end product application, transportation and mining costs, and local government regulations including zoning. The present major aggregate producers are most likely to develop these future resources and are in the best position to initiate comprehensive exploration programs.

Jan 1, 1987