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By T. K. McCranie, A. H. Tousley, A. H. Kackman, L. R. Gregory, A. Jr. McElrath, R. J. Krekel

In Roman and earlier civilizations the term cement was applied only to mixtures of lime, pozzolana, sand, water, etc., used as a mortar to bind larger stones. Today, portland cement, the subject of this section, can be briefly defined as follows: The product obtained by finely grinding clinker produced by calcining to incipient fusion (i.e., sintering) an intimate and properly proportioned mixture of argillaceous and calcareous materials. History The Egyptians were probably the first to join building stones with a mixture of sand and a cementitious material (i.e., having the property of or acting like cement). It is generally accepted that the cement used by the Egyptians on such structures as the Great Pyramid was calcined gypsum, Later the Greeks began to calcine limestone for use as a mortar with sand and water. Still later, broken brick and tile were added to the mortar to make the first concrete. The Greeks and Romans found that some sands produced mortars that were especially resistant to the action of water. Hydraulic (i.e., underwater-hardening) cements formed by combining slaked lime, water, and finely divided siliceous materials (e.g., pozzolana) possess superior strength and are capable of resisting the destructive action of water. They were used on the Pantheon, Colosseum, and other structures, some of which are still standing. In 1756 an English engineer, John Smeaton, found that an argillaceous limestone (i.e., containing an appreciable amount of clay as an impurity) produced a cement with greater resistance to the action of water. Despite Smeaton's findings, the use of the old mixture of lime and pozzolana long retained its popularity. Significant contributions in cement technology from England, Swe¬den, France, and Holland were made in the next 68 years. However, the invention of portland cement is generally credited to Joseph Aspdin, an Englishman, who in 1824 patented "portland cement." The name was selected because, on hardening, it resembled the natural building stone from the Isle of Portland, England. The patent features mixing ground limestone and argillaceous materials in appropriate proportions and calcining (i.e., expelling CO2 by roasting) the mixture. This superseded the existing practice of calcining naturally occurring argillaceous limestone. In about 1845 Isaac Johnson, also English, conducted experiments on proportioning the components and calcining them at higher tem¬peratures. His accomplishments are accepted as the beginning of the present-day portland cement industry. The U.S. cement industry originated with the demand for good hydraulic cement for building structures such as the Erie Canal, which was started in 1817. The first discovery of cement rock (i.e., naturally occurring argillaceous limestone with near-optimum ratios of calcium carbonate, alumina, and silica) near Fayetteville, N.Y., was made by C. White. In 1818, White secured a patent for the manufacture of natural cement from that deposit. Subsequently, several natural cement plants were built in 1830-1840 at Rosendale, N.Y., and in the Lehigh Valley of Pennsylvania. David Saylor, who participated in establishing a natural cement plant at Coplay, Pa., in 1850 discovered that a superior cement could be produced by calcining the rock at higher temperatures than hith¬erto. In 1871, Saylor obtained a patent and commenced the manufac¬ture of portland cement. By 1890, there were 17 portland cement plants in the United States with a combined annual output of 62,300 tons. At the end of 1970, the number of plants producing portland cement in the United States (including Puerto Rico) was 169, down from a maximum of 185 in 1967, with total annual shipments of 73.4 million tons valued at $1,298 million. (Fig. 1). In 1970, annual world production of hydraulic cement (which includes pozzolanic materials, hydraulic limes, natural cement, alumina cement, and portland cement, all of which are in current use in various parts of the world) reached 630 million tons (Table 1). Flowsheets of typical U.S. cement plants are shown in Chapter 13. Chemical Composition and Physical Properties It is generally accepted that portland cement clinker consists of a mixture of compounds or synthetic minerals. The four principal compounds are:

Jan 1, 1985

By G. Wayne Clough

During shallow tunneling in an urban environment, deposits or pockets of saturated soil or fractured, weathered rock are often encountered. Under such conditions, special problems with tunnel support, underpinning of structures adjacent to the tunneling, and/or flow of groundwater into the tunnel occur. If the ground conditions are appropriate, i.e., the soil is sandy, silty or fissured, or the rock is weathered or fractured, the solution to these tunneling problems may be chemical stabilization of the ground. Injection stabilization as an aid to tunneling is practiced extensively in Europe and its use is increasing in the U.S. The advantages of this technique over conventional solutions can be substantial in terms of cost and time, especially in mechanized tunnel boring where even a small zone of soft ground can bring the tunneling to a complete halt. In this paper European practice in using chemical ground stabilization for urban tunneling will be reviewed. In contrast to previous papers which have dealt with specific applications (1,5,8,9) this paper will broadly cover the subject, addressing itself to grouting techniques, costs, the reasons for grouting, philosophy of the users towards grouting, design of grouting solutions (or lack thereof) and methods of quality control. Differences with U.S. grouting practice are also noted. The information presented is derived from several visits to Europe, discussions with European and U.S. grouting personnel, and the available literature (2). The author's most recent European trip was made under the auspices of the U.S. Department of Transportation in conjunction with a research study being performed at Stanford University; a detailed report of this visit is available in reference (6).

Jan 1, 1976

By Manfred Hager

The ability of conveyor belts to transport large mass throughputs economically over previously unprepared ground has resulted in this system achieving great and extensive use. A significant component in this development is the conveyor belt itself. The development of high strength steel cord conveyor belts involves the optimising of splice design, the use of excellent rubber material especially in the splice, good craftsmanship during splice production and modern field vulcanisation equipment. The durability of a splice for belts of class St 3000 up to St 790 is expressed by the fatigue strength under dynamic stress. The results obtained with a test method and equipment developed by the University of Hannover indicate the present state of the art in this field. Belts of high nominal strength used on inclined long distance conveyors reach splice fatigue strengths of about 3000 N/mm. The design should only exploit approximately 50% of the fatigue strength determined in the tests as maximum operating stress. IMPORTANCE OF CONVEYOR BELT SYSTEMS The use of conveyor belt technology has expanded considerably over the last decades in the bulk goods transportation sector. Because of the favourable transport costs and the technology's adaptability to specific topographies, belt conveyors frequently represent the most economic solution. When flows of goods are large, as in German lignite mining operations, where masses of up to 40,000 t/h are given, it is the only technically feasible alternative. The use of belt conveyor systems allows the sensible distribution of the waste and the production of the coal (Hager, 1981). But this transportation alternative is also suitable when the masses involved are smaller. Whereas conveyor belts were only suitable for loose bulk goods up until a few years ago, today they are frequently also used for hard rock open cast mines, e.g. in the production of copper ore. Because of the favourable transport costs it is frequently still economic to use mobile crushers with a throughput of up to 10,000 t h 6 which breakdown the large blocks produced y explosives into transportable grain sizes. In many operations, trucks are only used to provide flexibility between the excavator and the crusher, i.e. over short routes, and the long, frequently steep, transport paths to the processing plant or to the spreader are undertaken by belt conveyors (Einenkel et al., 1992). The advantages of belt conveyors, i.e. to cope with large mass flows over inclines and down slopes of up to 1:4 have resulted in this system gaining very extensive use throughout the world. The conveyors can be installed over for the most part unprepared ground with suitable vertical radii, adapted to the locally available belt material. In the process individual conveyor belts with lengths of up to 15 km, and in the underground sector with lift heights of up to 1 km, have been built and operated. This great variety of application possibilities is complemented by a limited ability to pass through horizontal curves. This can be achieved with the help of design measures on the conveyor whilst bearing in mind the characteristics of the belt. STRESSING OF CONVEYOR BELTS IN DIFFERENT TYPES OF PLANT The advantage of conveyor belts compared with other systems is also to be found in the large available service-time window, i.e. the excellent reliability and the low costs of energy consumption and maintenance. It is in particular the reliability of a belt transport system which depends to a great degree on its main component, that is the belt - in all its different types and variations. For this reason, the belt is also given the greatest attention in the development of the individual components, because it is the belt which must be designed optimally with regard to various factors. The specific properties of the conveyor belt then assert a considerable influence on the design and sizing of most other components in a

Jan 1, 1993

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

INTRODUCTION The majority of ground water flow models have been developed for large computer systems. Most of these operate under a batch mode of operation, that is, the data are input with a deck of punched cards. The advent of micro- and minicomputers over the last few years opens the possibility of adapting many of these large flow programs to the smaller computers. In general, this adaptation will require more computer time to run the program, but in many cases if the small computer is in-house, the overall cost may be less than that required for a large computer. There are also a number of peripheral devices that can be used to make the data input much easier. As an example, the use of a digitizer with a finite element or finite difference mesh reduces the data input time considerably; the use of a graphics terminal for display of the mesh immediately after it is developed shows immediately whether there are errors in the location of node points. DEVELOPMENT OF AN INTERACTIVE SYSTEM The finite element program, UNSAT2, has been used over the past several years by the authors for several ground water flow problems (Bloomsburg, 1977; Bloomsburg and Wells, 1978; Zahl and Bloomsburg, 1980). In the past these programs have been run on an IBM 370/ 145 computer under the batch mode of operation. One difficulty that has been encountered has been in determining the length of time step that must be specified for operation of the program. The problem is that if the time steps are too large, the program becomes un- stable and operation ceases. If this happens in the middle of the problem, the only alternative is to reduce the length of time step and repeat the entire run. This problem has been alleviated in other pro- grams by using a subroutine that determines the time step automatically. If the solution becomes unstable, the time step is reduced automatically and that portion of the solution is repeated. With an interactive system this can be done by storing the output pressures and restarting the program with the new output pressures. Under the batch mode of operation, there is an option in UNSAT2 that allows restarting of the program from the last point of solution, but this must be specified when originally starting the program. The interactive computer system that we use currently consists of a PDP 1 1 /23 with hard disk drive, CRT terminal, plotter, graphics terminal, digitizer, and printer. The computer has 256 K bytes of core storage and the hard disk drive will accept a 5.2 M byte disk. To run a large program on a computer such as this, some overlaying must occur; that is, subroutines are stored on disk and put into core storage only when they are actually needed for running the program. UNSAT2 required very few changes to run on this computer but some changes were necessary to take advantage of the interactive capabilities of the system. One of these changes consists of the development of a program called DBUILD to build the data file for running the program. This program calls for each piece of data by prompt statement and gives instructions for using the digitizer. The data are then entered in completely unformated form. The node points must be entered in order and the coordinates are determined automatically by the digitizer. The pressure readings or initial conditions on each node point are entered through the terminal. All element information also is entered through the terminal. After DBUILD is run the mesh may be displayed immediately on the graphics scope whereupon any error in node point

Jan 1, 1986

By Gaither H. Kitchell

Introduction Fire assaying is an ancient method of determining the amount of gold and silver in geological materials. Today's commercial, mining, school, and private fire assaying laboratories still use this basic assaying method. One of the primary reagents used to collect gold and silver from the ore is litharge (PbO), or lead oxide. The Federal Occupational Safety and Health Administration's (OSHA) and the Nevada Division of Occupational Safety and Health's (OSH) have a lead standard. It states that a person exposed to airborne lead at concentrations greater than 50 µg/m3 (1.4 µg per cu ft) of air over an eight-hour period has gone beyond the permissible limit for health safety. It is not uncommon to get airborne lead concentrations of 1400 µg/m3 (40 µg per cu ft) in the flux/sampling area or 900 µg/m3 (25 µg per cu ft) around the gas/ electric furnaces during the fusion and cupelling phases. Therefore, any individual working in these areas for more than 30 days per year without proper protection is a likely candidate for lead poisoning. Lead In A Human Lead enters the body as a toxic substance by breathing it into the lungs and upper respiratory tract and by getting lead particles into the digestive system through ingestion. Smoking, eating, drinking, chewing tobacco and gum, or wiping a lead-contaminated hand across the mouth are all potential ways forgetting lead into the body. Lead taken in large quantities can actually kill a person in a matter of days. This is rare, however. The real hazard of lead overexposure is for the unsuspecting person who works in an unsafe lead environment for months or years. This individual can slowly receive damage, sometimes permanent, to the central nervous system, gastrointestinal tract, kidneys, blood, gums, and reproductive system. Safe Blood-Lead Level The Lead Standard states that the maximum safe blood-lead level is 40 µg/100 g (40 µg per 3.5 oz) of whole blood. It recommends 30 µg/100 g (30 µg per 3.5 oz) if a person is of child-bearing age. People not working around lead products normally have a blood level of less than 10 µg/100 g (10 µg per 3.5 oz). A person not adequately protected can exceed 40 µg/100 g (40 µg per 3.5 oz) in a few weeks by fire assaying only 100 samples per day per week. Studies show that some individuals have developed lead-related diseases with blood-lead levels less than 80 µg/m3 (2.3 µg per cu ft). Lead Poisoning Symptoms There is a long list of symptoms associated with lead poisoning. The most common are: a prolonged, sweet, metallic taste in the mouth; loss of appetite; anxiety and irritability; weakness and dizziness; nausea and tremors; and insomnia. Preventive Measures Based on the lead standard, a number of preventive measures can be taken to ensure that fire assay laboratory personnel frequently exposed to lead are protected against overexposure. First, the actual lead concentration within the assaying area should be determined. In Nevada, one can contact the Department of Industrial Relations, Division of Occupational Safety and Health for assistance in monitoring any laboratory. The agency will also recommend ways to protect against lead. This agency does the monitoring and provides an educational class on lead safety. This type of agency can be found in other states. Commercial consulting firms are also available, but they are expensive. Regardless of the technique used to determine lead concentrations, it is most likely that the average fire assay area exceeds the 50 µg/m3 (1.4 µg per cu ft) lead concentration limit. This is considered unsafe.

Jan 8, 1985

By Donald C. Moore

Introduction Asamera Minerals (US) Inc. and its joint venture partner, Breakwater Resources Ltd., discovered ore grade gold mineralization on their 20 km2 (5000 acre) Wenatchee, WA land position in February 1983. Due to the high grade nature of the discovery ore and the known reserves of ore in the "B Reef' and "B West" zones previously outlined by other companies, a decision was made to construct a mine/mill operation near the known ore occurrences. Further drilling in the discovery area quickly expanded known gold occurrences to more than 3.6 Mt (4 million st) with tentative in-place ore grade of 7 g/t (0.25 oz per st) and minor silver values. Based on existing knowledge of the ore body and the rapidly increasing ore reserve, a decision to build a 1.8-kt/d (2000-stpd) mine and mill complex was made in the second quarter of 1983. A schedule was devised to begin immediate mine development, shaft sinking, environmental and land use permitting, and mill and tailings dam construction (Fig. 1). Meeting the scheduled startup date, April 1, 1985, required a fast track schedule in all areas. To this end, Asamera purchased the Oracle Ridge Partners concentrator. This was an assemblage of new equipment designed for use as a copper concentrator in southern Arizona. The purchase contained all of the major mineral dressing equipment - crushers, screens, rod and ball mills, etc. and an engineering package. It did not include most of the other required items, such as buildings, conveyors, pipelines, tanks, and pumps. At the same time, core samples were sent to two independent process development laboratories for initial flowsheet development. Due to the refractory nature of the carbonaceous ore, cyanide leaching was not feasible. Flotation was selected as the concentration process. Further testing showed that autoclaving of the flotation concentrate followed by cyanidation would result in overall recovery of about 85% gold. A mine manager was hired to begin assembling an operations staff, hire an environmental consulting firm, and begin mine development. Environmental and land use concerns were major obstacles due to the mine's close proximity to a city of 20,000 people. These concerns had to be rapidly defined so as to mitigate any adverse impacts from and mining processing operations. Baseline data dealing with weather, air and water quality, and sound were measured before start of mine construction. Concentrator and flowsheet development remained static until October 1983 while definition drilling and mine development proceeded. In late October, a process engineer was hired to coordinate development of a process flowsheet, purchase the remainder of the concentrator equipment, prepare a concentrator construction contract, finalize concentrator detail engineering, and combine environmental and process requirements with a tailings dam design. Process development There were only 17 months remaining to mill start up from the hiring date of the process engineer. Therefore, the process flowsheet had to be finalized rapidly. To accomplish this, samples of drill core from the highest grade (and therefore potentially the most commercial) ore zones were sent to an outside metallurgical laboratory to confirm beneficiation tests on the flotation process. Test results again showed that flotation would provide about an 86% gold recovery. Therefore, all further testing was concentrated on flotation and autoclave/cyanidation of flotation concentrates. Focusing on a well known process such as flotation was important in accomplishing the rapid design and construction of the concentrator. If, during these next phases, we were continually changing design concepts, layout, and process flow, the mill startup would have been delayed many months. Once a process flowsheet is selected the process engineer must obtain the process criteria needed to design the beneficiation system. For example, it was known in early December that the Oracle Ridge rod and ball mills were too small to grind 1.8 kt/d (2000 stpd) of Wenatchee ore. A decision had to be made to purchase a large, used ball mill and convert the Oracle Ridge ball mill to a rod mill. The process engineer must be cognizant of the process criteria needed to size and select equipment. If not, the process engineer must use the professional services of the equipment manufacturing companies to review the requirements that the equipment is asked to perform. For the Wenatchee system, this resulted in the adaptation of a ball mill to a rod mill with a weight limit of grinding rods to protect the mill bearings and drive trains. When a decision is required, the process engineer has to present the facts and options in a manner that allows a rapid decision. This information must include costs, equipment availability, and effect on the construction schedule. At the Cannon mine, there were process development details that resulted in decisions similar to the ball mill purchase. These included an increased flotation residence time from eight to 25 minutes, an increased thickener area requirement, a high pressure tailings pumping system, and area constraints in plant layout. All of these decisions had to be timely and required assistance from manufacturers' service engineers, and knowledge of the alternate costs and effects on construction completion. Equipment procurement It was decided in early 1983 to build the ore milling facility with Oracle Ridge equipment, augmenting it with used equipment

Jan 2, 1989

Introduction Surface mining is a business, and the objective of most businesses is to make as much money as possible within certain responsible constraints. Businesses normally measure their potential to make money by estimating the discounted cash flow rate of return on new projects or the net present value for ongoing projects. These estimators of profitability are calculated using the following well-known equation: [ ] where i is the year, n is the total number of years, A, is the after tax net cash generation in year I, r is the discount rate and NPV is the net present value. In its simplest form, when the value of r makes the NPV equal to zero, the discounted cash flow-rate of return (DCF-ROR) is defined. The maximum economically recoverable reserve is identified at the point where the NPV or DCF-ROR is at a maximum. Depending on the measure used (DCF-ROR or discount rate) for calculating the NPV, the definition of recoverable reserve will be different for exactly the same deposit. Maximization of the measure of profitability and the economically recoverable reserves should not be separated from the mining method and sequence of exploiting the reserve. The surface mining methods may involve multiple-pass development of the resource, mining to the ultimate limit in one pass or a combination of the two methods with backfill. Multiple-pass development When planning a multiple-pass mining operation, it is important to determine the width of pit expansions that are appropriate for a specific situation. Production personnel typically want wide, open work areas where operations can be set up to run at peak efficiency. While this is understandable, the trade-off between highly efficient production operations and an overinvestment in stripping should be carefully weighed. As long as safety is not compromised, a few cents per bank cubic yard increase in operating cost due to lower productivity by working in more confined spaces may well be offset by a decrease in the overall short-term stripping requirements. A mini study on productivity and cost should help quantify this cost before deciding on minimum phasing requirements. The bigger challenge associated with establishing smaller work areas is securing operations and management buy-in to marginally lower productivity when reporting results. When cost reduction is the important management goal and everyone is focusing on unit costs, one sometimes loses sight of the goal of total cost management and, even more importantly, maximum profitability. The first pit should be defined by the first grouping of recoverable mineral that has the highest after-tax unit value and forms a feasible mining unit. The assumed mining of the first pit should be followed by the assumed mining of the second most profitable practical pit or pit expansion. If this process of grouping and sequencing is followed until the profitability approaches zero, the general sequence required to extract the maximum economically recoverable reserve is identified. The curve in Fig. 1 is constructed for a rate of production, R, and a quality management strategy, Q, by calculating the NPV for a series of sequences, with each sequence assuming that the operations are terminated on successively lower profit pits. These values are plotted to form the curve in Fig. 1. If the operation ceases on a pit that has too high of a profit, opportunity is lost, and the NPV is lower than the maximum. The NPV curve peaks and starts to decline before the zero-profit pit is assumed mined. This occurs because the money invested in advanced stripping would have made more money invested at the discount rate than by mining the next pit expansion. Therefore, the maximum economically recoverable reserve is defined by the pit limit corresponding to the maximum NPV. There may be an opportunity to refine the definition of the "ultimate pit" around the maximum-value point. The precise definition of the ultimate pit early in the mine life may or may not be important today, depending on how far off in the future these limits will be reached. By the time an operation must decide on its last increment of economically recoverable reserve. much of the

Jan 1, 2000

By William Paul Robinson

I would like to first summarize the relevant health information relating to the workers and general populations around uranium mining as a basis for understanding the hazards associated with these operations and their available remedies. This discussion will summarize the literature on Radon and its daughters, among the most toxic uranium decay products. Second, I will discuss waste handling in the uranium industry: past, present and future, discussing exploration, mining, and milling and the volumes of water and solid and gaseous wastes which are associated with these various phases of uranium operations. This paper will also include a discussion of the various past and present materials handling technologies and their responsiveness to the concern for uranium waste isolation. I will then conclude with a summary of recent modifications in the relevant standards as applied to uranium mining and milling. Joseph Wagoner, a research epidemiologist with the U.S. Public Health Service has summarized the literature on uranium mining and its health effects in his written testimony as a commissioned witness in British Columbia Royal Commission of Inquiry into Uranium Mining in January of 1980, entitled Uranium, The U.S. Experience: A Lesson in History.1 Briefly summarizinq his report, as early as 1546, miners of uranium bearing ores in Central Europe were reported to have unusually hiqh frequencies of fatal lung disease2. As early as 1879, the real nature of pulmonary disorders among these miners was first diagnosed to be malignant neoplasia or lung cancers. Of 655 miners dying between 1875 and 1912, 40% or 276 of these Central Europe miners died of lung cancer.4 By 1939, it was reported that lung cancer mortality rate for miners in these Central European mining districts was several times (9.7 per 1000) higher than that for similar populations in Vienna, Austria (0.34 per 1000). Thus, before uranium mining began in the U.S., there was a considerable body of information linking uranium mining to lung cancers. The U.S. experience can be summarized in Table 1 also from Wagoner's B.C. Testimony.1 This table shows that through 1978, 205 respiratory cancer deaths have been observed among white underground uranium miners in the

Jan 1, 1980

By S. C. Gambrell

The practice of roof bolting in world-wide mining activities is very well known, having been used extensively for over 50 years. The theory of roof bolting, however, is extremely complicated, and only recently have investigators attempted to explain the theory. Discussions of the mechanics involved usually originate with actual experience gained in the use of roof bolts. The use of roof bolting in mines has grown dramatically during the period after World War II. Other scientific techniques have emerged in mining research in the same period. Included in these techniques is the study of rock mechanics, which has now developed into an impressive research effort. Various means of analysis have been utilized in rock mechanics, most of which reduce rock to an isotropic, homogeneous medium that could be easily analyzed. Because of the inherent nature of rock, direct research methods frequently are not certain to provide satisfactory results, whereas indirect methods must idealize a given set of conditions in order to be useful. One of the indirect research methods is the photoelastic technique, the theory of which is widely known and needs no elaboration. This technique has been used extensively to examine the state of stress around underground openings of various shapes in certain types of strata, and under various configurations (1,2,3,4,5). Much of the theory used now in the design of underground openings is based on photoelastic evaluation. Unless rigorously applied, how- ever, the photoelastic technique is at best a qualitative means of evaluation. It is useful, however, for unusual examples which can be more or less easily analyzed to obtain the overall state of stress which could not be determined by other means. Little work has been done in the photoelastic analysis of rock bolting, mainly because direct research techniques have proven useful.

Jan 1, 1970

By I. H. Warren

Hydrometallurgy has a well-known and established role in the production of manganese metal (1) and manganese compounds, including battery grade manganese dioxide (2). Also, several leaching processes have been developed and applied at various times for the recovery of manganese from low grade ores (3). Recently, interest in the hydrometallurgy of manganese has been stimulated by the problems presented in the processing of sea nodules (4). Fumes and dusts from the ferro-manganese furnaces also offer a potential field for the application of manganese leaching techniques (5). In treating low grade manganese ores a major requirement of the leaching process is that it should dissolve manganese from a mixture of manganese and iron oxides. The most selective reagent for this purpose is ammonium carbamate applied after preliminary pyrometallurgical reduction of the ore (6). Since soluble carbamate complexes of the postulated type (7) are formed by both manganous and ferrous, but by neither manganic nor ferric, ions, selectivety of manganese leaching is achieved by control of the reduction operation to yield, essentially MnO and Fe304? A potential improvement to this two step process would be to simultaneously reduce and selectively leach the manganese in one hydrometallurgical operation. This possibility has been experimentally studied and the results are described below. During the preliminary study of the stability of solutions of ammonium, ferrous and manganous carbamates at elevated temperatures, it was observed, contrary to the observation of Welsh (8) at lower temperatures, that

Jan 1, 1971

By K. C. Watts

The Silver City mining region is located in Grant County, southwestern New Mexico (Fig. 1). This area is well known, having produced large tonnages of base and lesser amounts of precious metals over the past several years. Mining within the region continues today, as well as exploration activity, which seems to have been on the increase since the recent discovery of a large skarn-associated Zn-Pb-Cu replacement deposit near Pinos Altos, a mineral district within the region (McKnight and Fellows, 1978). Because the Silver City area contains a number of mineral districts and retains promise for additional deposits, as suggested by widespread hydrothermal alteration and favorable igneous rocks, it was selected as the focus of two independent geochemical studies, both seeking to enhance anomaly-to- background contrast in order to improve our abilities to detect blind metal deposits or subtle manifestations of near-surface occurrences. The first of these independent studies involved the systematic sampling of those secondary Fe-Mn oxides that occur within coatings, selvages, and fault fillings, and then the selective extraction of these oxides from the host-rock matrix by chemical methods. The second study sought to isolate clastic Fe-Mn-rich materials from the heavy fraction of stream alluvium. The ultimate source of the clastic Fe-Mn oxides is presumed to be the same coatings, selvages, and fault fillings sought by the direct sampling of joints, but has been transported as a result of lateral erosional processes.

Jan 1, 1981

By Gordon F. Revey

In underground construction and mining work, situations occasionally occur where hard rock or concrete must be excavated at locations near critical structures or utilities. Fearing damage from blasting, contractors and engineers often explore the viability of using mechanical methods or expansive concrete to facilitate the excavation work. However, for moderate to large excavations, the prolonged duration and high cost of using these methods is often impractical. In these situations, under the right conditions, controlled blasting and blast-management methods can be used to speedup the work and lessen costs. The process of determining whether or not blasting is practical is often complex, particularly when risks are high. In this paper, general methods of conducting blasting risk evaluations and outlined and considerations for blast design and risk management are reviewed. This paper also includes three blasting case histories demonstrating how these techniques were used at recent projects. In the first case, a cautious blasting program was developed to facilitate the removal of concrete and rock within a large section of an active mineshaft. The second case involved critical blasting to excavate a dry lake-tap connection between a tunnel and a shaft at the Southern Nevada Water Intake No. 2 project in Henderson, Nevada. The third case history describes how novel blasting methods were used to carefully drop a 2,721,500-kg (6,000,000-lb) concrete caisson into a1.2-m (4-ft) rock socket at the Detroit Outfall No. 2 project in Detroit, MI.

Jan 1, 2001

By Y. C. Kim, W. L. Meyer, W. Cai

Although the first published optimum pit limit design algorithm by Lerchs-Grossmann dates back to 1965, the majority of the open-pit mines in the US has not as yet adopted such a rigorous optimization algorithm. Instead, the floating cone approach initially developed by Kennecott Copper in 1961 is still used today for various reasons. However, shortcomings with the floating cone algorithm, particularly for precious metal deposits with erratic mineral distributions, have provided a renewed interest in optimizing algorithms. This paper compares the performance of the floating cone algorithm with the rigorous optimization algorithm of the Lerchs-Grossman's graph theory and the newly developed maximum network flow algorithm by Hanson (1986). The maximum flow algorithm is a modified version of the logic as initially advocated by Johnson in 1968. Comparisons are made for three different deposits in terms of the total profits generated and the computation times involved. The results of the comparison show that both algorithms produced higher total profitability pits than the floating cone in all three deposits, although the increase in profits never exceeded 1 % of the base design obtained using the floating cone. Between the graph theory and maximum flow, graph theory produced consistently higher profits. The current version of maximum flow, however, cannot be modified to produce the optimum pit. Nevertheless, the authors prefer the maximum flow over the graph theory, simply because it is easy to program as well as to explain.

Jan 1, 1987

By Herbert M. Mandel

INTRODUCTION Designers of major public works projects in the past have displayed varying degrees of environmental consciousness, and tunnel designers in particular have always had to deal with certain environmental considerations - ventilation, lighting, aesthetics - as matters of primary design. Today, however, with a widening sector of the public becoming environmentally aware, tunnels are treated as elements of whole transportation systems, and are subject to increasing environmental scrutiny. The Route H-3 tunnels in Hawaii are examples of the time, care, and detail involved in a major rock tunnel project. Project Description The Route. Interstate Route H-3 is a proposed 10-mile highway crossing the Koolau Mountain Range on the island of Oahu, Hawaii, providing a direct connection between the Pearl Harbor area and the windward part of the island in the vicinity of Kaneohe. It is designed as a dual highway with provision to accommodate mass transit. As shown on Figure 1, the project includes two sets of twin bore tunnels (through the Main Koolau Range and through Red Hill), three highway sections of which two are through largely undeveloped mountain valleys (South Halawa and Moanalua), a viaduct, and three interchanges. From the Halawa Interchange to the Kaneohe Interchange, the proposed highway consists of two 3-lane roadways separated by a median where the highway is at grade. One lane of each roadway will not be constructed initially but will be landscaped and reserved for mass

Jan 1, 1974

By David T. Swanson

The term "pressure tunnel" in the broad sense includes all hollow spaces, aligned along an axis and surrounded by rock, which are suitable for pressurization. In particular, this category includes: • horizontal or slightly inclined water pressure tunnels; • vertical or inclined pressure shafts; • pressure chambers or shafts of surge chambers; and, • compressed air storage tunnels. The main problems in the design and construction of pressure tunnels result less from their length or size than from the requirement for watertightness and, dependent upon this, reliability in operation and economy in construction. Where the lack of sufficient overburden does not permit the rock to accept the internal pressure in the tunnel, or where this pressure is so high that the watertightness is in doubt although the stability of the tunnel shell is not in question, a pressure tunnel must nevertheless be reinforced. Surge chambers, pressure tunnels and tailraces which are subjected to relatively low internal pressure, usually not much in excess of 1.5 MN/m2, are eminently suitable for construction in post-tensioned concrete. Pressure shafts, which as a consequence of the high water head are subject to a considerable maximum service pressure, are usually lined with a steel tube, a considerable proportion of the internal pressure being transmitted to the rock by backing concrete. TYPES OF LINING FOR PRESSURE TUNNELS In general, the type of lining will depend upon the nature of the rock. As a result, pressure tunnels usually are not uniform in con-

Jan 1, 1981

By J. Köppel, K. Kovári, Ch. Amstad

INTRODUCTION The measurement of displacements in soil or rock in the vicinity of underground openings provide valuable information for a better understanding of the behaviour of the structure. On the basis of systematic observations conclusions can be drawn concerning - safety of the opening in the various stages of excavation - effectiveness of the support measures - nature of the rock pressure to be expected - influence of the construction of the opening on neighbouring structures (settlements of foundations, underpassing other tunnels etc.) - comparison of theoretical predictions with the actual structural behaviour (1). Today there are several types of suitable instruments available that can be employed at reasonable costs for such investigations (2). The most important ones are the multiple point borehole extensometer, the inclinometer and the multiple point borehole deflectometer. As is well-known, the extensometer only measures the displacement component in the borehole axis, while the inclinometer and the deflectometer provide monitoring of rock mass movements normal to the borehole axis. In the following a new instrument is described, which has been given the name Extenso-Deflectometer ISETH. In this single piece of equipment certain features of the above-mentioned three types of borehole instruments are incorporated. it enables the measurement of all three components of the displacement vector at consecutive points along the borehole, the axis of which may have any direction in space. This device is portable like the inclinometer. A simplified version

Jan 1, 1979

By G. F. Friel, G. S. Morrow, J. C. Edwards

An experimental program in the U.S. Bureau of Mines' Safety Research Coal Mine evaluated the effects of dead end crosscuts on the transport time of CO along a mine entry. The presence of carbon monoxide is a characteristic of a fire in its incipient stage, and is significant in the determination of a fire location in a mine from the analysis of arrival times of pre-set concentrations at CO monitors. In this investigation two entries, with a total of ten crosscuts, connected by a room were instrumented with CO sensors as part of a real-time monitoring and data acquisition system. Controlled release of 1% CO was monitored as the contaminant traveled the entries for several different configurations of CO sensors and exhaust fan settings. In addition to the normal longitudinal dispersion of CO due to turbulent mixing at the front of the CO wave, the induced rotation of air established within a crosscut by the entry flow diluted the CO concentration through entrainment. It was determined experimentally for the entry and crosscut configuration considered that the transport time for a specified CO concentration was 27% greater than the predicted time based upon entry volume only, but 13% less than the predicted time based upon entry and total crosscut volume. The proper interpretation of the measured transport time of a specified contaminant concentration between sensors is important for mine safety in the isolation of a fire and plans for miner evacuation.

Jan 1, 1993

By Jeffrey P. Carstens

INTRODUCTION Thermally induced fracture of rock is one of the oldest rock breakage techniques known to man. Primitive underground miners used fires to heat and crack rocks as early as the third century B.C. (Ref. 1). History aside, however, present-day tunneling and mining interests are reconsidering the use of modern heating units for rock fracture for some very good reasons. One of the most important advantages of many heating techniques is the lack of direct contact between the heating device and the rock. This means a lower wear rate relative to mechanical cutters, and no large reactant force to be supplied by the tunneling machine. In addition to avoiding physical contact, some heating systems (e.g., electron-beam guns and lasers) can deliver energy to a rock face much more intensely than can be conveniently handled mechanically. Thus, heating power densities of the order of lob to 107 watts/in.2 and greater can be achieved with high-power lasers. These high power levels allow much higher instantaneous penetration rates than can be achieved by mechanical cutting. Another advantage that accrues to some thermal techniques, such as, for example, surface spalling, is that the thermal fracture process works particularly well in certain rock types (e.g., granite) that are particularly hard to cut mechanically. Other methods, involving high power density, are able to cut any rock type, and are relatively

Jan 1, 1997

By Arnaud Colson

1 Good morning ladies and gentlemen. Let me first thank you for welcoming me to this international congress. I?ll have the pleasure to explain to you how the Lafarge group works on its quarry rehabilitation. I will deal with the way quarry rehabilitation has affected all branches of the group trough a global approach. 2 Quarries often are a subject of controversy dividing those are for and those who are against. My purpose is rather to examine where factors of progress lay and how in a quarry such as Lafarge we can include quarry rehabilitation, which is so fundamental dimension of our process. 3 - 4 As these two photos of French quarry show taking into account the environment in the quarry industry has been a recent preoccupation. In less than a century rehabilitation has a dramatically evolved in a positive way. Consequently our group integrate an ecological dimension in its activity. It?s an ethical question as well as a personal view of our role as a responsible industrial group. 5 My presentation is built around 4 points: I?ll begin by explaining why it?s our responsibility to rehabilitate quarries. I?ll go on to fill you in on the method of the project. After that I?ll outline the expected results. Finally I?ll talk about the rest of the project and the way we can ensure this know how.

Jan 1, 2002

By Brian Burdett, Terry I. Mudder, Robert T. Richins, James A. Stevenson, Michael M. Botz, Alan L. Wilder

The first Cyanisorb process was constructed in 1991 at the Golden Cross Mine in Waihi, New Zealand to recover cyanide from CIL tailings. The Golden Cross operation consists of an open-pit and underground gold mine with a conventional crushing, grinding and CIL circuit with Merrill Crowe zinc precipitation for gold recovery. Prior to disposal in the tailings pond, slurry tailings are stripped of cyanide at a pH of approxi­mately 7.5 using the Cyanisorb process. Mill throughput is approximately 2,500 tons per day (TPD), and CIL slurry tailings average 37% solids with a specific gravity of 1.27. CIL tailings WAD cyanide average 150 mg/L. The Cyanisorb process recovers approximately 85% of tailings WAD cyanide, produc­ing a slurry effluent with approximately 20 mg/L WAD cyanide. A second Cyanisorb process was commissioned at the NERCO DeLamar silver mine in southwest Idaho in May, 1992. The ore milling and silver recovery operation consists of crushing, grinding, a CIP circuit and Merrill Crowe zinc precipi­tation for silver recovery. Cyanide recovery is from clarified tailings with WAD cyanide ranging from 3 50 mg/L to 600 mg/L, producing an effluent with less than 30 mg/L to 60 mg/L WAD cyanide. Due to the presence of zinc-cyanide complexes in mill tailings resulting from the Merrill Crowe process, cyanide is stripped at a pH of approximately 5.5.

Jan 1, 1995