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By E. Stamboliadis

Gravity separation techniques were used in the early years to extract native metal nuggets and flakes from placer gold deposits. This was relatively easy for visible metal grains. As the size of the grain becomes smaller traditional gravity separation techniques become inadequate to recover native gold. Cyanidation was invented as a solution to the problem. According to it, gold and other precious metals are dissolved in water by sodium cyanide. The result was a drastic increase of world gold production but at the same time environmental issues were raised. In Greece gold mining has become equivalent with the absolute destruction of the environment and is under extinction. In the last years a new type of gravity equipment have been produced which create a gravitational field up to 200 G and can separate gold from gangue minerals at sizes below 50 µm. The strong gravitational field is created by centrifugal forces and the separation occurs in a fluidized bed, formed by water flow. Accumulations of finely disseminated native gold are found in the sedimentary sequence of the Flysch of the Ionian Zone. The main gold minerals are gold and electrum present as free metal flakes and micro nuggets in the size range between 5-250 µm. Test work, using the Falcon SB40 centrifugal concentrator, at the Department of Mineral Resources Engineering at the Technical University of Crete, has shown that this particular type of gold ore can be processed by gravitational methods that require no reagents and are environmentally safe.

Jan 1, 2003

By S. M. Moore

Each year more than 400 coal miners are injured (fatal and non-fatal) by rock falling from between or around roof supports. Many of these injuries could be prevented by the installation of roof screen (wire mesh). However, many coal mines may be reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential saving in workers? compensation (WC) premiums that could be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by IL and KY were investigated in this study. Using actual data obtained from MSHA, national and state WC bodies and individual insurance companies (e.g. average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), hypothetical mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state?s actual WC rate-setting formulas, total savings in WC costs after a three-year period were determined. Across both states, savings in WC premiums ranged from 1.8% to 14.6% when injuries were reduced by 10% to 20%. Additionally, an economic analysis was performed for an existing mine consisting of 67 employees. For this mine, the annual cost of a roof screening program was estimated at $240,000. If the screening prevented 13 ?struck by? WC claims over a 3-year period at this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By Egon Pohl, Friedrich Steinhäusler, Werner Hofmann

INTRODUCTION Anticipated increase of nuclear fuel production in the future coincides with growing concern about the occupational health risk of miners from inhaled radon decay products. As a consequence it has been suggested to even lower presently permitted exposure levels (NIOSH, 1980) and implement more stringent control on measurement programmes. Compliance with these regulations requires large investments in new or modified ventilation systems as well as in increased expenditure for staff and instrumentation for health physics operational monitoring. Since both costs are directly related to the overall ore production costs, this can have far reaching implications with regards to the economic feasibility of certain mining operations and consequently reduce estimates of low-cost minable ore reserves. In addition there is increasing evidence that the radon problem is not limited to the nuclear fuel cycle only, but can also represent a significant hazard to non-uranium miners. A key component in the cost-effective implementation of legislative control measures is the monitoring system employed. The choice of system is decisive for the total costs of installation and maintenance, manpower requirements and accuracy of nuclide determination. In the following operational specifications are defined for monitors in mining and milling environments. Different types of available instrumentation are discussed with regard to their suitability for practical radiation protection in underground mines, open cut mines and mills. ATMOSPHERIC CHARACTERISTICS Underground mines Radon in the air of underground mining operations is exhaled from surrounding rock surfaces, crushed material and, to a lesser extent, from water seepage. Whilst in uranium mines radon releasing ore bodies are generally localized in distinct areas, radon sources in non-uranium mines can be very dispersed throughout the system of mine tunnels. The ventilation scheme used influences the absolute atmospheric level of radon as well as the equilibrium conditions between radon and its decay products (factor F). In uranium mines mechanical forced-air ventilation is normally the only way to achieve and maintain legally required nuclide levels. This causes the F-factor to be rather low, e.g. in French CFA-mines F is of the order of 0.2 (Francois, Pradel, Zettwoog, 1975). At the same time the fraction of unattached radon decay products (fp) can increase due to the high air velocities employed. However, it is possible to find non-uranium mines with either natural draught ventilation only or assisted on demand by forced air ventilation during special operations or climatic conditions. Thereby the F-factor is more dependent on seasonal changes of temperature differences between outdoors and mine atmosphere and work routines. As a result F can cover a wide range from 0.02 to 0.95 (Steinhäusler, 1976). The use of filters or electro-precipitators in mine ventilation systems can modify the atmospheric characteristics twofold as it generally decreases the content of radon decay products, but at the same time increases the content of the unattached fraction fp . Average concentration levels of radon decay products are mostly lower in mechanically ventilated non-uranium mines than in equally ventilated uranium mines and are below 0.3 Working Level (UNSCEAR, 1977). However, some working places in non-uranium mines, specially with only natural draught ventilation can occassionally approach maximum permissible levels as defined for uranium mines (Strong, Laidlaw, O&apos;Riordan, 1975; Snihs, 1976; Sciocchetti, Scacco, Clemente, 1981). Open pit- and surface mines Radionuclide levels of radon decay products in the atmosphere of these mines are mostly too low to represent a significant inhalation hazard for miners, ranging typically from 0.03 to 0.1 Working Level (Steinhäusler, 1976). However, personnel using airpurifying respirators or working in cabins ventilated with filtered air can be exposed to a radon atmosphere with low value for the F-factor (F [<] 0.1) and high fp-value up to 80 % (Leach and Lokan, 1979). Mills Atmospheric radon concentration in crushing, grinding, drying and packing sections depends on the radium 226 content of the ore, ore storage methods and ventilation system employed. Providing adequate ventilation ([>] 2 air changes per hour) and control of dust production radon and its decay products represent only a minor problem (Saconney, 1979). MONITORING OF OCCUPATIONAL EXPOSURE Objectives Operational monitoring of the working places provides information on: - confirmation of appropriate control of the routine mining methods employed - indication of abnormal conditions. Although this type of monitoring enables the location

Jan 1, 1981

By A. Papandreou

Industrial wastes nowadays constitute one of the most serious threats for the future of our planet. Researchers around the world have very often proposed many kind of solutions in the management problem of industrial effluents. However in most cases suggested solutions are either too expensive to be applied in industrial scale, or they just transfer the problem (e.g. precipitation creates dangerous sludge) and so they create an even greater problem and threat. In the present study the exploitation of valuable hydraulic and adsorption properties of fly ash is studied. Fly ash can easily and cheaply be shaped into stable pellets. After ageing for a sort period of time, these structures can be used as adsorbents for the removal of heavy metals from industrial effluents. Experiments in aqueous solutions have shown excellent adsorption behavior concerning the removal of cadmium, copper, lead, chromium (III) and zinc. The advantage of this technology, except low cost, is that after saturation the pellets can be easily removed from the liquid phase and stabilized in concrete structures without any problem of desorption or leaching. Moreover in these structures a significant amount of cement in the concrete can be replaced by raw fly ash. (Ashcrete as it is called). The utilization of an industrial byproduct as fly ash in the form of pellets as adsorption means and their stabilization with concrete (or ashcrete) combines the useful properties of a solid waste in the treatment of liquid effluents and constitutes a green low cost technology.

Jan 1, 2006

By Priscilla P. Nelson, Francis L. C. Fong

Introduction Fracture mechanics is the study of the behavior of cracks and their modes of propagation. During cracking, energy is consumed in creating new surface area. The more energy a material absorbs, the higher is its fracture resistance. This fracture energy is provided either by work done externally, by the release of elastic strain energy within the cracked body, or by a combination of these two effects. Fracture resistance is measured in terms of either the stress intensity factor, K, or the energy release rate (also referred to as crack driving force), G. For linearly elastic materials, these two quantities are related by the following equation: [ ] where [u] is Poisson&apos;s ratio and E is the modulus of elasticity. Both K and G are direct functions of the applied force and the crack length. The process of rock cutting with full-face tunnel boring machines (TBMs) involves the indentation of a rock surface by rolling disc tools. Excavation occurs as fractures, extending from the locus of chip indentation, propagate between concentric disc grooves and liberate rock chips from the intact rock face. Since disc cutting is achieved by a fracture process, it is logical to anticipate that differences in cuttability of a variety of rocks should be related to the variation of fracture material properties among the rocks. Such a relation has been demonstrated between the performance of full-face TBMs and the critical crack driving force for excavation in relatively high strength, brittle, low porosity, isotropic rock (Nelson et al., 1985). The objective of this paper is to report on the results of a continued investigation of the relationship between fracture material properties and the effectiveness of disc cutting. Additional reliable data on full-face TBM performance is difficult to obtain, so that the results of disc linear cutter tests performed by the Transport and Road Research Laboratory (TRRL) (U.K.) are utilized in this study. Fracture Material Property Evaluation If a crack tip is subjected to a perfectly plain strain condition, many materials will behave in a brittle manner, and the material resistance to fracture is constant during crack growth. This resistance is referred to as the plain strain fracture toughness, believed to be a material property. The standard testing procedure for this parameter is given in the American Society for Testing and Materials (ASTM) test designation E399. To ensure plain strain conditions, acceptable specimen geometries are specified. Unless these geometric specifications are adhered to, the measured fracture resistance is regarded as the apparent fracture toughness, and a subscript Q is used, as [ ]. The methodology adopted by ASTM E399 is within the framework of Linear Elastic Fracture Mechanics (LEFM). This approach assumes that material resistance against cracking is developed instantaneously, without initial stable or subcritical crack extension. Prior to crack advancement in rock, however, much of the crack driving energy is absorbed in a localized volume in front of the crack tip, referred to as the plasticity or microcracking zone. This capability for energy absorption substantially increases material resistance, and a crack may be arrested unless an increased crack driving force is applied. The LEFM or plain strain assumptions would underestimate the actual fracture resistance of such a material. Such a conservative approximation may be desirable in the design of structures against fatigue or brittle failure. In mining or tunneling, however, an underestimation of crack resistance can lead to the selection of underrated excavation equipment and a decrease in production efficiency. The development of fracture resistance in rock is best described by the material resistance curve, or R-curve, which is a plot of fracture resistance as a function of crack extension. Typically, peak resistance is only reached after a finite amount of subcritical crack growth. The

Jan 1, 1986

By QUANWEN PU, Yonggang Wei, Bo Li

The spent catalyst is one of solid waste resources, which contains precious metals with considerable recovery value. However, due to the different compositions of catalyst, the recovery process is difficult. Most of the catalysts are supported by alumina which has good alkali-soluble properties, while platinum is insoluble in alkali. Platinum is enriched by pressurized alkaline solution which is a kind of hydrometallurgical process. The novelty of this study is the separation of alumina and platinum from the spent catalyst and the enrichment of platinum in the solid phase. In this study, a leaching process using NaOH as leaching agent under high temperature and pressure was employed. The effects of alkali concentration, leaching temperature, leaching time, and liquid–solid ratio on leaching efficiency were investigated. The optimum process conditions were determined as follows: alkali concentration of 240 g/L, temperature of 200℃, pressure of 1.20 MPa, time of 1 h, liquid–solid ratio of 5:1. It was concluded that the leaching rate of alumina can reach 99.69%. The content of platinum increased from 0.1749 to 21.12%, the loss of platinum was 3.39%, and the enrichment ratio of platinum increased by 120.82 times. The leaching kinetics of alumina was studied based on the shrinkage core model. The activation energy of alumina leaching is 19.53 kJ/mol, and the leaching process is controlled by chemical reaction. Pressurized alkali solution can not only efficiently leachate alumina from the spent catalyst and enrich platinum, but also reduce the reaction time and increase reaction rate.

Feb 11, 2022

By M. Fornaro

A small size quarrying activity (3 different quarries with an overall production in blocks less than 10.000 m3/year) has been exploiting for a long time the ?Tonalite? deposit of Adamello-Presanella. In 1967, given the outstanding natural beauty of the area, a natural park has been established with the aim of preserving environmental quality and, at the same time, of providing a sustainable development to the area. Because of the potential impacts of a mining activity in such a context, a specific study has been recently carried out in order to define if and how the quarries could co-exist with the park, beyond the ?administrative results? of a classic environmental impact assessment. The paper deals with the process that drove to the conclusion that the activities are not ?a priori? incompatible with the aims of the park. In the opinion of the authors, a responsible quarry management, which must take inspiration from the care given to environment by traditional alpine productive activities (e.g. the use of water, of wood, and of the stone itself), and take advantage of modern technologies in terms of emissions and safety, can be considered not only compatible with the protected area, but even able to positively contribute to the protection of the land and environment (Barisone et al., 2001). A series of specific prescriptions, which have been adopted by the Park Authorities after this study, are also discussed in the paper: only the definition of suitable spatial and temporal limits, the adoption of effective technical practices and a strict control are considered to be able to drive this ?exceptional? activity towards the final goal of real sustainability.

Jan 1, 2007

By J. C. Johnson

The problem of estimating pillar capacity in coal mines has motivated development of a variety of empirical equations. Most of these equations have arisen from experience and a very limited set of measurements that have, necessarily, been focused on particular regions or coal fields. Their applicability is often limited to a specific range of pillar sizes. This is of particular concern in the design of large gateroad pillars and small barrier pillars between panels. The result has been a variety of equations and equation coefficients that can calculate load estimates to size pillars. Inclusion of these equations in more sophisticated loading models, such as boundary element methods (e.g. LaModel and MULSIM) has been hampered by absence of a function that describes specific variation of coal strength within the pillar. A singular exception, to our knowledge, is a function corresponding to the Bieniawski pillar strength equation. Because of this, it is the only coal strength criteria inherently available in most boundary element models. This paper presents a general method for deriving local coal strength, as a function of distance from the pillar rib, for an arbitrary pillar strength equation. Integration of this function provides the strength of pillar regions, or cells, needed for boundary element models. The method is demonstrated for the Bieniawski, Maleki and Holland-Gaddy equations. The Maleki and Holland-Gaddy equations contrast strongly with Bieniawski as strengths are extrapolated to large pillars and barriers. This development provides a tool for incorporating any of the proposed empirical equations into a boundary element model, thereby allowing the analyst to choose the most appropriate equation for a particular region, coal seam and/or pillar size.

Feb 23, 2014

By C. Clark, J. Seymour, D. Benton

"National Institute for Occupational Safety and Health NIOSH is conducting research on jackleg use and related accidents in underground metal mines. This paper provides an analysis and overview of jackleg drill usage, accidents, operational characteristics, and alternatives, based on information from injury reports, legacy research, stakeholder input, and published literature. Jackleg drills are involved in more ground fall accidents in underground metal mines than any other drill. Jackleg-related injuries are most prevalent at the face in the course of installing initial ground support. Alternatives to jacklegs for drilling and bolting under incomplete support in narrow underground openings are not yet available. Small mechanized bolting equipment needs to be developed to address jackleg-related accidents and improve safety at mines where jacklegs are being used. JACKLEG DRILL Description The term “jackleg drill” is a generic name for a class of handheld, rotary, percussion rock drills that are equipped with an attached reaction leg to provide stability and thrust for the drill (Figure 1). All of the drill and leg controls are located near the operator’s handle at the rear of the head portion of the drill. Jacklegs are classified according to the bore size of their pneumatic pistons which normally range from 6.03 to 8.26 cm (2.375 to 3.25 in) [Kurt, 1982; Clark, 2012]. Jacklegs typically weigh 460 N (103 lbs), but their weight can approach 535 N (120 lbs) after they are fitted for operation with drill steel and air and water hoses [McKibbin and Clark, 2001]. In underground mines, jackleg drills are commonly used to drill 3.8-cm (1.5-in) diameter holes 1.83 m (6 ft) deep in rock. Application The jackleg drill is primarily used in underground mines when the size of the ore body, the ground conditions, or the mining method do not permit large openings to be mined using mechanized equipment (e.g., narrow vein mining). Jacklegs are used to drill blast-holes for explosives and also holes for installing ground support. These versatile drills are used in nearly all phases of mining including exploration, development, production, and maintenance [Kurt, 1982; Williams et al., 2007; Chen and McKinnon, 2012]. The jackleg is used to drill holes in rock using drill bits attached to hollow steel bars called drill steels (rotary percussion drilling). After the hole is drilled, the jackleg can be used to install different types of ground support components, depending on the ground conditions and the specific ground support measures that are needed. For example, the jackleg can be used to install expansion-anchor rockbolts or resin-grouted rockbolts using the drill’s rotary-motion feature. On the other hand, the jackleg can also be used to install friction-type Split-Set™ rockbolts using the drill’s percussion (hammer-motion) feature. As shown in Figure 1, various washer-like bearing plates can be installed along with these bolts to provide additional ground support and to help secure surface control products, such as wire mesh and shotcrete that are in turn used to retain small, loose material between the bolts [Clark, 2012]."

Jan 1, 2016

By J. D. Megeath

In the October 1988 issue of MINING ENGINEERING, J.G. Wargo&apos;s paper discusses the problem of measuring and analyzing permeability, porosity, and fracturing in its relation to in situ leaching of disseminated gold deposits. This paper made me suspect that a good deal of information and analysis may have been done on this in reverse in the petroleum industry. Some years ago, I was involved with the US Geological Survey in a project studying the Devonian shale of the eastern United States. The Devonian shale contains gas deposits that have the characteristic of slowly seeping gas over a very long time period. To understand the rate of seepage, considerable study and analysis was done on the interrelationship of porosity, permeability, and fracturing. The purpose was to find out how and why the gas could escape in the varying rates found in wells throughout the formation. It strikes me that this is the mirror image of the problem faced in attempting to understand how a cyanide solution will travel into a deposit. To possibly invent our own personal wheel a little faster, I would suggest a look at the work done on the Devonian shale. Some of it is contained in "Estimates of Unconventional Natural-Gas Resources of the Devonian Shale," US Geological Survey Open-file Report 82-474, 1982, and "Additional Data for the Estimates of Unconventional Natural-Gas Resources of the Devonian Shale," Open-file Report 83-676, 1983. It is entirely possible that some of this previous work could shorten the time required to make in situ leaching of disseminated gold deposits a profitable technique.

Jan 1, 1990

By J. D. Megeath

In the October 1988 issue of MINING ENGINEERING, J.G. Wargo&apos;s paper discusses the problem of measuring and analyzing permeability, porosity, and fracturing in its relation to in situ leaching of disseminated gold deposits. This paper made me suspect that a good deal of information and analysis may have been done on this in reverse in the petroleum industry. Some years ago, I was involved with the US Geological Survey in a project studying the Devonian shale of the eastern United States. The Devonian shale contains gas deposits that have the characteristic of slowly seeping gas over a very long time period. To understand the rate of seepage, considerable study and analysis was done on the interrelationship of poros¬ity, permeability, and fracturing. The purpose was to find out how and why the gas could escape in the varying rates found in wells throughout the formation. It strikes me that this is the mirror image of the problem faced in attempting to understand how a cyanide solution will travel into a deposit. To possibly invent our own personal wheel a little faster, I would suggest a look at the work done on the Devonian shale. Some of it is contained in "Estimates of Unconventional Natural-Gas Resources of the Devonian Shale," US Geologi¬cal Survey Open-file Report 82-474, 1982, and "Additional Data for the Estimates of Unconventional Natural-Gas Re¬sources of the Devonian Shale," Open-file Report 83-676, 1983. It is entirely possible that some of this previous work could shorten the time required to make in situ leaching of dissemi¬nated gold deposits a profitable technique.

Jan 1, 1989

By Fernando H. Muge

The first application of dynamic pro¬gramming to the problem of grade control and scheduling was by Muge et al (1979) and Ribeiro (1982) who applied it to the simple, two-dimensional case of longitudinal open stoping using two sublevels. In this paper the formulation of the problem is extended to cover the general case applied to a cut-and-fill mining method. The criterion for optimization is the total deviation of the average, grades of mining blocks from the overall average panel grade (or the average grade of a set of mining panels or the mine processing working point). This means to schedule the operation by removing a certain number of mining blocks at a time, subject to the geometric (mining) constraints, and such that the total deviation of average block grades from overall average grade is minimum. The method can be applied to complex orebodies with several ore characteristics and the deviation of each one of these characteristics can be further constrained by specifying maximum values. The method can also be generalised in order to take into account any necessary geometric constraints used to define the mining method. In this paper a case study is presented concerning the mine planning of a complex sulphide orebody located in the South of Portugal. As input for the definition of the mining sequence, a geostatistical numerical model of the orebody was built based on a previous ore typology. The short term planning is used to assess specific mining sequences and the production scheduling is performed on the basis of an objective function (e.g. minimization of grade fluctuation).

Jan 1, 1989

By John Marsden, Mike Botz

A variety of modeling techniques can be utilized to forecast metal production at heap leaching operations. These approaches reflect a wide range of complexity, flexibility, time to implement, cost, and accuracy. For many operators, a spreadsheet-based modeling technique is attractive since the calculations are directly accessible, models can often be developed by site staff, and the results are generally easy to extract and interpret. The authors have developed a spreadsheet-based modeling technique that provides a high degree of flexibility, while still considering detailed operating information for ore properties, leach kinetics, scaleup factors, lift height, and in-heap metal inventories. The technique involves establishing kinetic leach curves for each ore type to characterize metal extractions, followed by application of separate metal recovery curves to define the rate at which leached metals exit the heap in pregnant solution, taking into account delays due to the in-heap solution inventory. At any point in time, the difference between total metal extracted and total metal recovered is equal to the in-heap inventory of the leached metal. This modeling technique is particularly useful for larger multi-lift heaps where delays in metal recovery are appreciable due to in-heap solution holdup. The technique is also applicable to sites with multiple heaps, leach cycles, and/or metal recovery plants (e.g., carbon columns for gold; solvent extraction for copper). This paper describes the modeling technique, including input data required and methods for defining kinetic leach curves and metal recovery curves. The technique is compared against an alternative spreadsheet technique using a simplified in-heap solution inventory calculation to forecast metal production, which has proven to be effective at a number of sites.

By John Marsden, Mike Botz

A variety of modeling techniques can be utilized to forecast metal production at heap leaching operations. These approaches reflect a wide range of complexity, flexibility, time to implement, cost, and accuracy. For many operators, a spreadsheet-based modeling technique is attractive since the calculations are directly accessible, models can often be developed by site staff, and the results are generally easy to extract and interpret. The authors have developed a spreadsheet-based modeling technique that provides a high degree of flexibility, while still considering detailed operating information for ore properties, leach kinetics, scaleup factors, lift height, and in-heap metal inventories. The technique involves establishing kinetic leach curves for each ore type to characterize metal extractions, followed by application of separate metal recovery curves to define the rate at which leached metals exit the heap in pregnant solution, taking into account delays due to the in-heap solution inventory. At any point in time, the difference between total metal extracted and total metal recovered is equal to the in-heap inventory of the leached metal. This modeling technique is particularly useful for larger multi-lift heaps where delays in metal recovery are appreciable due to in-heap solution holdup. The technique is also applicable to sites with multiple heaps, leach cycles, and/or metal recovery plants (e.g., carbon columns for gold; solvent extraction for copper). This paper describes the modeling technique, including input data required and methods for defining kinetic leach curves and metal recovery curves. The technique is compared against an alternative spreadsheet technique using a simplified in-heap solution inventory calculation to forecast metal production, which has proven to be effective at a number of sites.

By John Marsden, Mike Botz

A variety of modeling techniques can be utilized to forecast metal production at heap leaching operations. These approaches reflect a wide range of complexity, flexibility, time to implement, cost, and accuracy. For many operators, a spreadsheet-based modeling technique is attractive since the calculations are directly accessible, models can often be developed by site staff, and the results are generally easy to extract and interpret. The authors have developed a spreadsheet-based modeling technique that provides a high degree of flexibility, while still considering detailed operating information for ore properties, leach kinetics, scaleup factors, lift height, and in-heap metal inventories. The technique involves establishing kinetic leach curves for each ore type to characterize metal extractions, followed by application of separate metal recovery curves to define the rate at which leached metals exit the heap in pregnant solution, taking into account delays due to the in-heap solution inventory. At any point in time, the difference between total metal extracted and total metal recovered is equal to the in-heap inventory of the leached metal. This modeling technique is particularly useful for larger multi-lift heaps where delays in metal recovery are appreciable due to in-heap solution holdup. The technique is also applicable to sites with multiple heaps, leach cycles, and/or metal recovery plants (e.g., carbon columns for gold; solvent extraction for copper). This paper describes the modeling technique, including input data required and methods for defining kinetic leach curves and metal recovery curves. The technique is compared against an alternative spreadsheet technique using a simplified in-heap solution inventory calculation to forecast metal production, which has proven to be effective at a number of sites.

Sep 3, 2019

By Carsten Rühlemann, Thomas Kuhn, Cornelia Kriete

"Any economic calculation with respect to potential submarine mining activities requires reliable analytical data on the contents of valuable metals in these ores, including trace metals such as rare earth elements as well as Ti, Mo, V, Zr and Li. Due to the very special matrix of ferromanganese nodules (30% manganese oxides, 10-20% iron oxides, silicates; see Kuhn et al., this volume), their geochemical analyses pose a real challenge. Available ferromanganese nodules reference materials, e.g., NOD-P1 and A1, JMn-1, GBW07249, SDO-4 to -7 and NM2388, generally have only a few certified values for trace elements. Published data for these reference materials, collected in the GeoReM database (http://georem.mpch-mainz.gwdg.de/) are either rare for high technology metals or widely scattered for certain elements, especially the PGEs. Since most of these materials were released in the 1990s, many analytical procedures have seen significant advancement.In order to facilitate analytical quality assurance during exploration and potential future mining activities, the German Federal Institute of Geosciences and Mineral Resources (BGR) has contracted the Central Geological Laboratory of Mongolia (CGL) to produce a certified reference material for nodule analytics. Certified reference materials provide the highest level of reliability in geo-analytical measurements currently available and the CGL has a long-term experience in producing geological reference materials. Recently, the CGL was accredited as a reference material producer according to ISO Guide 34 by the American organization ACLASS (Assured Calibration and Laboratory Accreditation Select Services). Furthermore, it has been accredited as a testing laboratory according to ISO 17025."

Sep 1, 2014

By Timothy J. Haddon

For the mining industry in general and gold mining in particular, the past few years have been difficult. Economic conditions, coupled with growing regulatory restrictions, have led to increasingly important reasons for managing assets wisely. From a public company perspective, such asset management makes sense because it ultimately provides for an increase in shareholder wealth, and that is why we are in business. Regardless of whether a company is public or not, assets must be managed effectively to remain a player in today&apos;s competitive environment. So, as the chief executive officer of a gold mining company, how do we manage our assets to ensure that we are satisfying our obligations to our customer-the Amax Gold shareholder? The gold industry has been under siege. We have just completed the third year with average prices below $12.85/ g ($400/oz). When finding, and/or acquisition costs and development costs to operating costs are added, there isn&apos;t much of a profit margin left. What this environment provides, however, is opportunities for growth. In the gold business, you either grow or you die. Today, most North American gold producers have, at most, a 10-year mine life. We are all scrambling to add reserves so we do not become extinct. As prices remain low, opportunities abound for increasing diminishing reserves. This is the time for the opportunistic company to take advantage of market conditions and that is precisely what Amax Gold has done. There are six components to running a successful gold business: • finding reserves through exploration, acquisition or joint venture, • financing the project, • permitting and constructing the project, • operating the project, • cash management and • communication.

Jan 1, 1992

By John W. Harshbarger

The Upper Santa Cruz Groundwater Basin lies in the drainage area of the Santa Cruz River in Arizona, and extends upstream from the community of Rillito to the international boundary, see Figure 1. The principal water use area lies between and includes the cities of Tucson and Nogales. A subsurface geologic barrier near Rillito provides a reasonable dividing line between the groundwaters of the Upper and Lower Santa Cruz Basins. The Upper Basin contains the Tucson and Sahuarita-Continental critical ground- water areas. The Tucson Water Use Area encompasses the Tucson metropolitan area and extends from the northern boundary of the Basin south to the boundary between the Tucson and Sahuarita- Continental critical groundwater areas near San Xavier. The San Xavier Use Area is the San Xavier Indian Reservation. The north and south boundaries of the Continental Use Area coincide with the limits of the Sahuarita-Continental critical groundwater area. The first study of groundwater in this valley was initiated by Dr. G. E. P. Smith in 1905 and the results were published in 1910, followed by a series of bulletins of the Arizona Agricultural Experiment Station, University of Arizona, covering the general field of groundwater development and utilization. Beginning in 1939, the Ground Water Branch of the United States Geologic Survey, in cooperation with the Arizona State Land Department and the Arizona Water Commission (AWC), have published a series of bulletins concerning groundwater occurrence and utilization in various areas of the state. In 1975 and 1977, the AWC published the "Arizona State Water Plan" (1), comprising a detailed inventory of water availability and use in the State of Arizona, and a study of alternative futures.

Jan 1, 1979

By S. M. Moore

Each year more than 400 coal miners are injured (fatally and nonfatally) by rock falling from between or around roof supports. Many of these injuries can be prevented by the installation of roof screen. However, many coal mines are reluctant to use roof screen because of the added cost. The goal of this study was to determine the potential savings in workers&apos; compensation (W C) premiums that can be achieved due to a reduction in rock fall injuries after roof screening. The WC rate-setting methods utilized by Illinois and Kentucky were investigated in this study. Using data obtained from the Mine Safety and Health Administration, national and state WC bodies and individual insurance companies (e.g., average cost per injury, loss cost rate, number of injuries per year, number of injuries preventable each year with roof screening), baseline mines (representing two mine sizes: 67 and 150 employees) were constructed with realistic ranges for estimates of injuries and WC premium costs. Using each state&apos;s actual WC rate-setting formulas, a sensitivity analysis was conducted to determine the total savings in WC costs after a three-year period. Annual savings in WC premiums ranged from $41,000 to $326,000 for the 67-person mine, and $96,000 to $843,000 for the 150-person mine. An economic analysis of the cost of a roof screening program at a 67-person mine was also conducted. The annual cost of a roof screening program here was estimated at $240,000. At this mine, the reduction in WC premiums alone could pay for the entire screening program.

Jan 1, 2010

By Markus Thewes, Mario Galli

"INTRODUCTIONProject-related studies showed that lab tests are a helpful tool for the planning of EPB tunneling and soil conditioning in cohesionless soils. This type of ground commonly represents the extended application range for EPB shields, which demands for a treatment with additives. Considering soil conditioning with foam, so far, approaches exist for the investigation of the foam quality and of the conditioning behavior of soils (e.g. (EFNARC GUIDELINES 2001 2003)), but not any standards for testing procedures related to soil conditioning. State-of-the-art research on the topic is and was conducted for example in Bochum (Germany) cf. (Budach 2012; Maidl 1995), in Lyon (France) cf. (Quebaud et al. 1998), in Oxford and Cambridge (Great Britain) cf. (Borghi 2006; Peña Duarte 2007), in Delft (Netherlands) cf. (Bezuijen and Schaminée 2001), in Turin (Italy) cf. (Borio 2010; Peila et al. 2013; Vinai 2006), and in Aachen (Germany) cf. (Vennekötter 2012). Additionally, applied science is reported by the tunneling practice and suppliers of conditioning agents. Although the research focus in all these and further studies varied, similar test methods and test series were applied. A standardization of essential properties and testing methods (setup and procedures) however did not take place. The results from laboratory testing can only function as index values for practice because in the studies, production and application of the conditioning agents mainly takes place under atmospheric conditions. Nevertheless, the tests enable basic understanding of feasibility as well as processes and interactions going on during soil conditioning. For recommendations for standardized and reproducible testing see (Budach 2012; Budach and Thewes 2015).The most extensive investigations on the flow behavior of soil-foam mixtures have been performed by using the slump test from concrete engineering, cf. (DIN EN 12350-2 2009-08), which originally is applied on evaluating the workability of fresh concrete. Although this test does not describe the actual flow behavior, it rather represents an index on the workability, very much experience exists with this test method. Besides others, the workability – or flow behavior – of the EPB support medium certainly is a major factor of influence on the properties of the material in the excavation chamber and thus, for effective tunneling with EPB shields, cf. (Galli and Thewes 2014). For a summary of various investigations on the workability of conditioned soils using the slump test and resulting recommended slump ranges see (Budach 2012; Vinai 2006). Thereof, slumps between 10 and 20 cm have been reported as suitable workability for EPB applications. However, there is actually not much known about the real flow behavior of soil-foam mixtures related to rheological properties such as the viscosity (?) or potential yield stresses (t0). These physical values are necessary for a realistic determination of face pressure, for the design of required torques, and for an evaluation of the material flow at different locations in the excavation chamber, in the screw conveyor or on the conveyor belt. To generate a linear pressure distribution at the tunnel face, as it is assumed a priori, a low-viscosity fluid is favorable, whereas in the screw conveyor, a high-viscous fluid is needed for a controlled pressure decrease. On the conveyor belt, yield stress plays an important role preventing deliquescence and spilling from the belt. Attempts in characterizing the actual flowing of soil-foam mixtures have so far only been made by few researchers, the application of rheometry is rare."

Jan 1, 2016