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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 This superseded the existing practice of calcining naturally occurring argillaceous limestone.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 Richard L. Stein

INTRODUCTION Increasing a miner's chance of surviving a disaster requires advance planning for emergencies, adequate training of personnel, and the provision of proper survival equipment. It is extremely important that fires and toxic gases be detected early, and that warning systems such as stench, audible and visual alarms, and communications systems be provided to allow timely escape. However, situations do occur where timely warning is not provided and immediate escape is not possible, regardless of the precautions that are taken. Under such circumstances, a combination of new per¬sonal protective devices and old techniques can provide protection that is adequate to save a miner's life. SELF-RESCUE EQUIPMENT Self-rescuers are devices which provide a short-term supply of respirable air under the potentially lethal con¬ditions following a mine disaster. Functioning either as filtration devices for the elimination of carbon monoxide (CO) or as oxygen supplies, self-rescuers allow an endangered miner a limited amount of time in which to effect an escape or reach a place of temporary safety. Carbon-Monoxide Self-Rescuers For almost 50 years, filtration self-rescuers have protected miners' lives following explosions or fires. Designed to eliminate carbon monoxide from inhaled air, the first of these devices was approved in the 1920s and has evolved into the two units that are approved today-the Mine Safety Appliance (MSA) W-65 and the Draeger 810. Both of these units rely on the cata¬lytic conversion of very toxic carbon monoxide to rela¬tively safe levels of carbon dioxide (CO,). As shown in Figs. 1 and 2, both of these filtration self-rescuers operate in the same manner. On inhalation, the mine air passes through both coarse and fine filters that remove the dust, preventing the dust from coating the chemical beds or entering the miner's mouth. The air then passes over a drying agent that removes water vapor; this is required to prevent the water vapor from poisoning the catalyst. Subsequently, the dried air passes over the Hopcalite catalyst, where carbon monoxide is converted to carbon dioxide. Then, the air passes through a heat exchanger that reduces the temperature of the inhaled air. Finally, the air is inhaled by the user. Air exhaled by the user passes through the heat ex¬changer and exits the device through an expiratory valve. Both units of this type are subject to certain limita¬tions: 1) The units do not protect the user against oxygen¬deficient air; when air is inhaled through this type of self-rescuer, the device only removes the carbon mon¬oxide. If the mine air contains less than 15% oxygen, anoxia is inevitable. Symptoms of anoxia include dizzi¬ness, shortness of breath, quickened pulse, and deeper and more rapid respiration while the victim is at rest. During heavy exertion such as would be expected dur¬ ing escape efforts, a 15% oxygen level can cause loss of consciousness. 2) The units do not protect against excessive levels of carbon dioxide. Available data on carbon monoxide and carbon dioxide concentrations following explosions or fires are limited, but there are indications that the carbon monoxide concentration can rise to 2% and the carbon dioxide concentration can reach 5 or 6% of the mine air (by volume). In some situations, the in¬haled air could contain up to 6 or 7% carbon dioxide. At a carbon dioxide concentration higher than 2%, breathing patterns can be affected adversely. At a con¬centration of 6 or 7%, the effects include severe respira¬tory distress, with unconsciousness resulting from exer¬tion such as that needed for escape activities. 3) The units do not protect against high inhalation temperatures. The catalytic oxidation of carbon mon¬oxide to carbon dioxide is an exothermic reaction that evolves a large amount of heat. As the carbon monoxide concentration increases, the temperature of the inhaled air also increases. Available data show that above 1.5% carbon monoxide, the air inhaled through some self¬rescuers can be as high as 90°C (194°F). At those

Jan 1, 1982

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By A. Hayatdavoudi, E. Apande

Introduction Troublesome shales have plagued the petroleum industry for more than 50 years. Unstable boreholes have been experienced when drilled into shale formations; sane have been studied almost at the beginning of rotary drilling. A little progress has been made since then, but much is still to be learned. No simple solution exists so far, but good drilling practices combined with good mud practices have been helpful. It is well known that shales in general have a strong affinity for water, and when water is adsorbed, the stability of the shale section of a well is reduced; cohesive strength will be re- bed, properties will be altered and shale will generally expand and/or crumble. The amount of water a shale will adsorb depends on the characteristics of its clay mineral and the ionic concentration of surrounding fluid. Fluid loss and the resulting hydration of an unstable formation can lead to serious downhole problems. The rate of fluid loss depends greatly on the permeability of the deposited cake. In impermeable formations, however, there is no loss of fluid into the formation. As such, shale, because of its very low porosity and almost no permeability, could be termed impermeable. Yet there is strong evidence that shale surfaces are wetted by an invasion of fluid into the formation through microfractures and jointed planes. It is this wetting that causes the shales to swell and slough into the hole, resulting in problems such as increased mud volume and treating costs, bridges and fill-up, stuck pipe, overgauge hole, poor cement jobs and increased cement requirements, and logging and completion difficulties. Drilling operations have been defeated by severe shale problems, resulting in & hole being plugged and abandoned after several weeks of drilling. Shales are we1l known as being difficult and costly to drill. Numerous fishing jobs have plagued drilling operations, and the cost of reaming tight holes ad weighting up the drilling mud to control shale sloughing has been high. The concept of controlling water loss to stop sloughing was introduced in the 1920's. It was known-then that calcium in fresh water re- strains clay £ran excessive swelling (theory of osmosis) . This reinforced the belief that calcium-based muds such as lime, gypsum and calcium chloride should minimize hydration of clays. Lime muds were then introduced for this purpose and others in the 1930-1940 decade1. They were used for shale inhibition but were not effective. In an attempt to improve them, a special calcium-type mud beam popular during the 1950's. Hole enlargement continued and they too were considered ineffective. During & early 1960qs, papers were published claiming that shale swelling could be prevented by using high concentrations of chemical thinner called Chrane Lignosulfonate1. This began the era of highly treated lignosulfote muds. Unfortunately, the hole enlargement problem continued. During & mid-1960’s, oil based muds were used for the purpose of inhibition of shales. Not until 1966 did Mondshine2 suggest that the water phase be saturated with calcium chloride. The cost of using oil based mud was considered by most operators not worth the additional hole stability. The potassium chloride-polymer muds were also used in many fields in the United States and western Canada. The use of these muds has been claimed to prevent clay swelling by inhibition and encapsulation in sane areas of western Canada. They are also expensive and their ueneral application is still experimental. To reduce filtrate loss when sloughing shale becomes a problem during drilling is a reasonable practice. This study will investigate the use of cationic organic additives to mud to form a protective barrier between the shale - which are primarily composed of clays - and water that will keep water from entering the microfractures and will also prevent them from hydrating. In so doing, an understanding of modern clay chemistry and & dispersive properties of cationic additives in water based systems will be our most valuable tools. This study will also present and solve a mathematical model for simulating cation transport phenomena in fractured shales using the Framdlich equilibrium adsorption isotherm. The studies will analyze wellbore failure and stability as affected by cationic organic material.

Jan 1, 1986

By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By G. Kraemer, J. A. Le Gac, P. R. ZETTWOOG

I. INTRODUCTION During the course of our activities in assisting mining companies, we have had access to the monitoring results for personnel from a large number of uranium mines. The results differ greatly from one mine to another. One of our objectives has been to discover the origin of these differences. It is evident that they are largely due to variations in the geological context, to mining methods, and to the organization of the work ; but we also found that the rigor with which measures are implemented to prevent the personnel from being exposed to radiations is also a cause. In order to advise mining operators effectively, we have asked ourselves the following question Given the dosimetric results of a mining site, how can it be known - if occupational exposure has been reduced to the lowest possible level (dose optimization principle recommended by the ICRP) ; - if the exposures were justified, which would not necessarily be the case following a questionable choice of the mining method, for instance, or a lack of efficiency in the maintenance and repairs of ventilation apparatus or even an excess of radiation protection. It is necessary to establish some criteria, while taking into consideration the specific conditions for each mine, in order to determine whether a mining company is adequately implementing radiation protection procedures. This need led us to attempt a modelization of the occupational exposures of uranium miners ; the preliminary phase is presented here. Although voluntarily still very basic, this model makes it possible to demonstrate the role of certain dynamic or passive variables. Moreover, the model presents the concept of specific irradiation of a mining site equal to the collective dose received per ton of uranium metal supplied to the uranium mill. The specific irradiation can therefore be used to indicate the effectiveness of radiation protection procedures at a given mining site. This model can be used for - previsional exposure studies based on the use of data gathered at each site, making it possible to compare various work methods and to determine prevention means ; - qualification of "radiation protection" procedures at a mining site ; - detection of unjustified exposures ; - research of ways to reduce inevitable exposures. 2. DETRIMENT TO BE ASSOCIATED WITH A GIVEN EXPOSURE DISTRIBUTION Respecting exposure limits ensures that delayed stochastic effects are the only health risks to workers. If the hypothesis of linearity in the dose-effect relationship is applied, the detriment to a group of workers is proportional to the sum of exposure for all the workers, or the collective exposure usually expressed as man x rem. On this point the linearity of effects hypothesis makes it possible to consider that two different standard deviation distributions having the same mean value are equivalent. It would be possible, then, to reduce the standard deviation by changing the distribution of workers at the various working places without changing the mean exposure value. In other words, without taking into consideration the practical implications, individual limits can be respected by rotating the personnel or by reducing the individual working period which would be compensated by increasing the number of workers. The collective dose would remain the same, which would justify the use of the concept of collective exposure. 3. "RADIATION PROTECTION" QUALITY INDEX AT VARIOUS MINING SITES Each mining site is responsible for its own collective dose, but it is also responsible for producing a certain amount of uranium metal in the ore sent to the uranium mill. These two quantities must be brought together in order to establish the "radiation protection" quality index of a mining site. This quality index, called-specific irradiation (Ir) is expressed in rem.ton -1 for external irradiation, in mJoule.ton-1 for inhaled [a]-energy, and in Ci.ton -1 for inhaled radon *. The specific irradiation represents the health hazard which must be associated to the extraction of one ton of uranium metal. Figures la and lb present a system of coordinates for which the axis of the abscisses is graduated in tons of U metal and the axis of the ordinates is graduated in units of collective exposures. In the EEC, the regulation is based on the measurement of the amounts of radon 222 inhaled.

Jan 1, 1981

By Peter J. Huck, Yoginder P. Chugh, M. Jennings

INTRODUCTION Large areas of the United States coal reserves have been undermined by room-and-pillar mining over the past century. These abandoned mines generally cause subsidence of the ground surface many decades after mining. Reports from England (1) where there is a longer history of room-and-pillar mining indicate that mines abandoned two centuries ago may still be causing subsidence, and causing surface destruction to urban areas or farm lands that have been developed in the interim. The devastation wrought by subsidence, both in terms of property loss and societal disruption is enormous. The authors discuss here the methods by which subsidence in abandoned mines may be controlled. Two distinctly different subsidence control technologies have been developed in the U.S. for abandoned coal mines. These are referred to as (1) Point Sup- port methods, and (2) Areal Backfilling. As the names imply, point support methods are used to protect individual structures, or even individual foundation elements of structures, whereas areas backfilling techniques may be applied to protect entire neighborhoods or urban districts hundreds of acres in extent. Both point support methods and areal backfilling have specific sub-categories of techniques that may be applied under different conditions as appropriate. POINT SUPPORT METHODS Point support methods are characteristically employed by civil engineers chartered with the design of a new structure or the protection of an existing structure. The subsidence control measures are only a part of a larger project, the safety of which must be assured to a degree of near certainty. The costs of point support methods involve a large number of bore-holes and the use of expensive materials in relatively small quantities. There exist several dozen point support methods, which operate either to support the underground cavities, controlling subsidence, or to isolate the structure from the effects of subsidence if it should occur. The latter methods usually involve deep foundations through the mine opening supporting the structure on the underlying strata. Our concern will be with the former method to support underground cavities to prevent subsidence. Gravel Columns and Associated Methods Gravel columns may be placed in the mine opening using the methods shown in [Figure 1]. Boreholes are drilled through the mine strata, and gravel poured down the borehole to form a pile on the mine floor. When the tip of the pile contacts the mine roof, it is rodded down to spread the pile, and permit additional gravel to be placed. The objective is to place as much gravel as possible in the mine opening, and to achieve firm contact with the mine roof. The gravel pile acts to reduce subsidence by three mechanisms. The filling of a significant fraction of the open mine volume reduces the amount of potential subsidence simply by killing open volume within the mine. In effect, the extraction ratio has been reduced by introducing the volume of gravel. The toe of the gravel pile abuts against nearby pillars, providing lateral support and protection against pillar deterioration. In effect, the height of the pillars has been slightly reduced by burying their lower portions. Finally, some direct support is provided to the roof strata, reducing the roof span and enhancing stability. The gravel column method may be used in a variety of applications, and with materials other than gravel. For example, gravel columns may be placed in a close line around the perimeter of the building to form a continuous gravel wall. The interior of the site may then be completely filled with sand, slurry or other material. Alternatively, a lean concrete may be used in place of gravel, so that the piles achieve significant structural capacity. The grout column, shown in [Figure 2], begins with the installation of a conventional gravel column which is placed around a grout pipe that extends through the borehole into the mine opening. When the gravel column is in place, portland cement grout is injected in stages. At each stage, sufficient grout is injected to fill the interstices in the gravel to build a column perhaps two meters in diameter and a meter tall. The group pipe is then raised one meter, and after the previously injected grout has set, the next stage of grout is injected. The process builds a column of grouted gravel to the

Jan 1, 1982

By W. J. Schlitt

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

Jan 1, 1998

By J. C. Yingling, J. W. Leonard

Introduction The absence of coal-face particle management heavily penalizes the transportation of coal from initial loading to final consumption. The penalties include dust problems, significantly reduced mine-loading-cycle productivity, mine-belt spillage, excessively high coal-preparation costs, chute blockages and dangerous pulverizer blockages at the final point of utilization. Fine particles commonly cause environmental and economic problems. It is well known that these fines can cause safety and environmental dust problems. But it is not well understood that these fines can also swell broken coal to a point where 5% to 15% more time and capacity must be used to deliver the same tonnage. In this paper, methods and rewards for reducing and/or managing fines at the mine face are discussed. Computer-based loading-cycle model productivity estimates, viewed from a new perspective, are made on the basis of material volume rather than on the long-established, and frequently misleading, basis of tonnage. It is typically the volume of broken material being transported that defines the capacity of a given transportation system, while the corresponding tonnages are merely a reflection of the specific material densities. Published evidence suggests that the swelling of broken coal can be decreased very significantly using small quantities of certain nonfrothing chemicals, which are added to mine-face spray water, and by employing improved mine-face breakage practices. In a future paper, the effects on transportation productivity beyond the coal mine will be discussed. The precursor to the work presented in this paper, involving the bulk density improvement for broken coal and the subsequent production gains for underground coal mines, was earlier presented in Leonard and Newman (1989). In the past, this topic has been studied and practiced only in byproduct coking in the steel industry. However, a potential exists for an increase in coal-industry productivity by improving the bulk density of coal to yield a subsequent reduction in delivered cost. This can occur with breakage, handling and treatment methods resulting in the loading of greater quantities of coal in fixed volumetric capacity haulage units such as mine cars, shuttle cars and scoops. Laboratory-based experiments to achieve an increase in productivity by increasing coal bulk density were discussed in Leonard, Paradkar and Groppo (1992). Chemical techniques using small quantities of commercially available reagents (surfactants) resulted in about a 13% to 15 % increase in bulk density, which was thought to produce a proportional increase in the productivity of a mine, together with a subsequent reduction in cost. The idea is to mix the reagents with the water that is used to spray coal during mining. In this paper, the impact of bulk density improvements on production rates is presented. Increases in production ranging from 60% to 88% of the bulk density increases are projected. This analysis was performed for atypical continuous-miner section. In the following sections, discussion and results of the analysis are presented. Discussion An analysis was performed to ascertain the impact of bulk density improvements on face-production rates for a typical continuous-miner section. Figure 1 illustrates the section layout and cut sequence. This layout and sequence is identical to the case described in King and Suboleski (1991). As can be seen, the section uses five entries and 12.2-m cuts that are taken by a remotely controlled continuous miner. The seam height is 1.5 m and two shuttle cars (5.7 t nominal capacity) are employed for haulage from the miner to the section feeder, which, throughout the cut sequence, is positioned as illustrated in Fig. 1. The simulation model was coded in the SIMAN simulation language. The major impacts of increased bulk density improvements on such a production system are as follows: •Shuttle-car payloads, in terms of the mass of coal transported per haul cycle, are increased proportionally to the increase in bulk density that results from the application of surfactant. •Shuttle-car discharge times should remain largely unchanged, because they are determined by the volume of material that is discharged, rather than the mass, and this volume does not change.

Jan 1, 1996

By Yi Luo

Longwall coal mines are frequently located in steep mountainous areas with narrow valleys. In order to protect surrounding watersheds and residential structures, support areas are left under the valleys. However, the indirect subsidence effects to the residential structures due to the potential topography effects can be a serious problem. The present study examines a longwall mine in southern West Virginia located in an area with an average slope of about 30°, or 58%, and with an overburden depth that varies from 49 to 376 m (160 to 1,235 ft). A comprehensive investigation was carried out to assess the subsidence and topography effects, as well as other ground control issues at this site. Surface and subsurface movements and deformation were predicted and their direct effects on nearby streams and structures were assessed. The potential and indirect subsidence influences of surface topography effects were assessed. The techniques and results of this investigation are presented in this paper. It shows that the direct and indirect subsidence effects on the streams and residential structures should be insignificant. The post-mining observations agreed well with this study.

Jan 1, 2010

By F. Warnaars

Historically, this area produced gold and silver from many well-known vein deposits, like Tayoltita, Topia, and San Jose de Gracia. Modern exploration identified three categories of deposits: 1) low grade porphyry golds in shallow level dacite porphyries; 2) batholith-related deposits; and 3) those formed in volcanic sequences of Late Cretaceous to Late Tertiary. At least four porphyry golds have been identified, of which Metates is best known. These are formed by and within dacite porphyries, associated hydrothermal breccias and sometimes within surrounding metasediments. Their ages are from mid-Cretaceous to Early Tertiary. Precious metal deposits associated with batholiths occur as veins, contact haloes in metasediments or volcanics, skarns, tectonic breccias, hydrothermal tourmaline breccias, dacite dikes or as disseminated or stockwork deposits such as La Fortuna. Several new gold discoveries were made in older Early Tertiary Volcanic Sequences and in the younger Tertiary volcanics within prominent structures exceeding 3 to 5 km in lengths. Some of these are open-pittable but most have to be mined by underground methods.

Jan 1, 1999

By John M. Hird

Texasgu1f's Lee Creek Mine in Eastern North Carolina is usin8 a combination Dredge-Dragline method for mining deep phosphate ore at a rate of 8.5 million tons per year. The matrix is thick, uniform in grade, and is processed into agricultural products. As in many other good mineral deposits, there are tough obstacles to overcome in recovering the ore. The deposit is in an area with a land surface of about ten feet above sea level, is close to tidewater, is overlain with 90 to 130 feet of unconsolidated overburden, and is underlain by an artesian aquifer with over 140 feet of natural hydrostatic head at the bottom of the bed, The upper 40 feet of overburden will not support stripping equipment and is unstable when cast as spoil. Barriers are emplaced to prevent flooding from hurricane tides. The ground water problems have been successfully controlled by pumping 60 million gallons per day from the aquifer. Only recently have the severe ground problems been overcome by this unusual Dredge-Dragline combination of mining equipment.

Jan 1, 1978

By Alejandro Sanz, Juan Luis Magro, David Mazeyrie

The Hampton Roads Bridge-Tunnel Expansion Project (HRBT), initiated by the Virginia Department of Transportation (VDOT), will add two new 1.5-mile TBM bored tunnels to the two existing immersed-tube tunnels between artificial islands, part of the connection between Norfolk and Hampton under the Hampton Roads Harbor in Virginia, USA. The new bored tunnels will be built by Hampton Roads Connector Partners (HRCP), a joint venture of Dragados USA, Vinci Construction, Flatiron Constructors and Dodin Campenon Bernard. This paper covers technical discussions that led HRCP to select a Variable Density TBM to address the challenges of this tunnel project, including but not limited to (i) the diameter of the bore: a 46 feet (14.04 m) diameter machine that constitutes the second-largest TBM ever used in the United States; (ii) shallow ground-cover under the water table; (iii) challenging and varying geology along the alignment, with mixed-face conditions and different types of ground modification zones; and (iv) the high TBM operating pressures required—all these and other challenges in a sensitive marine environment.

Jun 13, 2021

By A. Platts, K. Leung, P. Rohner, P. Munro, B. Coulter

The Hilton Concentrator treats a laminated stratiform silver- lead-zinc ore to produce separate lead, zinc and bulk concentrates. Commissioning commenced in December 1989. Grinding circuit performance is above design with a mean flotation feed sizing of 80% passing 65 microns at 800 000 tonnes / year throughput. Grinding is done in two stages with an open circuit low diameter : length ratio SAG mill followed by a closed circuit ball mill. SAG mill control philosophy is to maintain a fixed feedrate for stability of flotation and to control mill charge weight to a setpoint with feed water. It is planned to increase throughput to 1 million tones / year in the second half of 1991. Proposed milling circuit modifications include close circuiting the SAG mill by recycling a portion of the ball mill cyclone underflow, increasing the SAG mill speed from 73% to 79% of critical and offloading of lower than feed grade -20mm /+7mm SAG mill ore scats. This paper will describe current mill operating practice with operating data and key mill developments since the beginning of operations.

Jan 1, 1992

By Louis W. Cope

INTRODUCTION Various methods for sampling placer deposits have been mentioned by Wojcik in Chapter 2. Here, a brief description and comparison of these methods will be given as related to bulk versus drill sampling. The advantages and disadvantages inherent to each will also be given. The means of sampling can be grouped into bulk or drill samples. As a generality, drill samples over a five-foot length will run from I/,, to 1 /, cubic yard, depending on the diameter of the drive shoe on the drill casing. Bulk samples can be expected to be in the range from one to many cubic yards each. SAMPLE METHODS Bulk Bulldozer-dug trenches and dozer or front-end loader pits are two methods to excavate large bulk samples. Smaller bulk samples can be obtained by digging small dimension shafts through the deposit, by hand or caisson-like drills. It is not recommended to split samples by alternate bucket-loads or by other means, as the effect of a nugget falling into one or the other fraction will seriously alter the sample.

Jan 1, 1992

By Annan Cook

Cyprus Metals' Sierrita mine is located in the Pima Mining District, Pima County, Arizona (Fig. 1), which is one of the largest known porphyry copper districts. The ore body is part of a large mineralized system that includes Cyprus' adjoining Esperanza mine. The copper and molybdenum grades being mined at Sierrita make it one of the two lowest grade porphyry copper mines in the world Mineral production from the district was first recorded in 1876. Copper from replacement deposits in carbonate rocks and copper¬-lead-zinc from veins in volcanic rocks were mined from several deposits. However, it was not until 1957 that the district commenced to become a major copper producer with the beginning of production from the Pima mine, which was the first open pit operation. Since that time, seven other open pit mines have come into production. Because of the variable copper prices and high operating costs, not all of the mines are operating today. A map of certain outcrops, plutons, and open pits, as well as older underground mining operations, in the Pima district is shown in Fig. 2.

Jan 1, 1991

By T. Weimer

Gasification of carbonaceous energy carriers with simultaneousCO2 removal is a new and interesting alternative e.g. ZECA [1],HyPr-Ring [2], AER [3]. In the ZECA process lime supported Hydrogen generation from Methane is a basic process unit and the Hypr-RING process investigates Hydrogen generation from hard coal with addition of lime. Here the main objective is a further development of the AER process. High energetic efficiencies for gasification in the range of 90 % can be achieved if the humidity of the energy carrier is already sufficient for steam reforming reaction. The aim of the overall process is upgrading coals and biomass yielding into three valuable products: • A fuel gas ready for power generation • A CO2gas stream containing >95%, ready for transportation to disposal or chemical fixation • A pre-calcined feed for cement clinkering consisting ofCaO and coal ash Thus power generation and gasification are independent. A promising industrial application is the combination of gasification and cement production. Then only one additional reactor is required and the precalcination mass flows have to be increased.

Jan 1, 2004

By Leonhard Schmid

INTRODUCTION In the region of Zurich several of the Swiss Interstate Highways meet and have to be linked to form a continuous network. The subalpine glacial topography with hills, ridges and lakes requires the construction of many tunnels (Figure 1). The Milchbuck Tunnel forms part of the inner arterial highway system of the city of Zurich. It passes underneath the ridge separating the Limmat and Glatt valleys. In its final configuration the tunnel will consist of two parallel three-lane tunnels (Figure 2). For the time being only one tunnel is constructed. This tunnel will be operated with two outbound (climbing) lanes and a single inbound (descending) lane (Figure 3). The owner, the Canton of Zurich, awarded the contract for the construction to a joint venture with the following three contractors: Locher & Cie AG (leader), Prader AG, A. Brunner's Erben, all located in Zurich. The actual freezing was performed by subcontractor Philipp Holzmann AG, Mannheim, West Germany. Electrowatt Engineering Services Ltd, Zurich, was design engineer and performed the construction supervision.

Jan 1, 1981

By Bruce E. Holbein, Dennis K. Kidby, Ann L. Huber

World gold production figures and the processes used for gold production indicate a high dependence on both cyanide leaching and carbon adsorption metallurgies. Problems associated with the use of carbon in the treatment of refractory ores as well as economic and environmental considerations associated with the use of cyanide point to the advantages that can be realized through the use of more efficient gold adsorption media and with cyanide recovery. We have developed a process termed AurosaveTM-1 which provides a number of advantages through the use of a new synthetic robust gold adsorption medium, termed Vitrokele TM V-911. This process allows further recovery of gold from a variety of ores and allows for a simplified process for use not involving thermal regeneration or high pressure elution. We have also developed a cyanide recovery process termed Cyanosave ™ which uses Vitrokele TM V-911 and allows efficient recovery of cyanides from decants or tailings and re-use of recovered cyanide. This obviates the high chemical expenses for cyanide destruction where environmental issues require such water and tailings management.

Jan 1, 1998

By S. J. Knoll, D. F. Scott, T. J. Williams

Highly stressed rock in stopes continues to be a primary safety risk for miners in underground mines because it can result in failures of ground that lead to both injuries and death. Spokane Research Laboratory personnel investigated electromagnetic (EM) emissions in a deep under-ground mine in an effort to determine if these emissions could be used as indicators of impending catastrophic ground failure. Results suggest that (1) there is no increase in the number of EM emissions prior to recorded seismic activity, (2) some EM signals are generated during blasting, (3) interference from mine electrical sources mask seismic-generated EM signals, (4) EM emissions do not give enough warning (compared to seismic monitoring) to permit miners to leave a stope, (5) the distance an EM signal can travel in the rock is between 17.7 and 39.6 m (58 and 130 ft), and (6) current data acquisition systems do not differentiate between EM signals generated from seismic activity and random mine electrical noise. These results preclude monitoring EM emissions as precursors of impending catastrophic ground failure.

Jan 1, 2004