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By Alan Bye

Sandsloot open pit is situated on the northern limb of the Bushveld Complex, 250 km north east of Johannesburg, South Africa. Sandsloot is currently the largest open pit platinum mine in the world and is one of six potential open pits in the area. The geotechnical strength properties of the Platreef have posed significant mining and processing challenges to the PPRust operation since inception in 1992. In order to optimise the whole business as opposed to isolated cost centres a mine to mill initiative was embarked upon in 1997. Initial work focussed on ore characterisation and an entire suite of data was collected for each rock type. This included mineralogy, geochemistry, geotechnical, blastability indices as well as metallurgical strength data. The second phase involved the clear definition of customer requirements, namely the optimum performance criteria for the plant and mining departments. This involved extensive field trials and data capture to measure the interaction and influence of the mine to mill value chain. Once the design targets were defined a 3D geotechnical model was constructed to ensure that the targets were consistently achieved. The model is very similar to an ore reserve model however it is populated with geotechnical, metallurgical and blast index data. Through software queries the appropriate blast powder factor for each mining block can be obtained thereby ensuring the correct fragmentation is blasted for the plant and mining operation. The initiative has delivered significant results showing improvements across the mine to mill value chain. These include increased loading rates, reduced electricity and crushing consumables as well as higher plant throughput. Additionally, the initiative has benefited both Brownfield and Greenfield projects on the PPRust operation.

Jan 1, 2005

By Jair Carlos Koppe, João Felipe C. L. Costa, Luis Eduardo de Souza

United Nations (UN) has been leading an international committee to define the guidelines in reporting resources and reserves. The system proposed by UN is very similar to the guidelines currently in use in Australia known as JORC guidelines (Joint Organisation Reserves Committee). The JORC guidelines show major differences in relation to the Brazilian system current in practice. The important aspects differing JORC from the current Brazilian guidelines for reporting reserves is the required sampling spacing. The JORC system requires the points of observation to be closer spaced than the Brazilian system for any class of resources. All classification systems share some common aspects in terms of defining the classes of resources based on distance separating samples and on the degree of confidence or accuracy associated with the results reported. Despite of being very clear in terms of stating sample distances, all the systems of classification do not provide clear definitions on how confidence limits should be calculated. This paper proposes a geostatistical framework to measure the uncertainty associated with the reported tonnages and grades. Results demonstrate that the error associated with an estimate can be approximated using multiple simulated deposits to define confidence limits. The methodology is illustrated in a large Brazilian coal deposit and proved to be appropriate to address the issue of quantifying the uncertainty necessary to be incorporated in the systems of reserves classification.

Jan 1, 2002

By R. A. Metz, George S. Barnett

The Equis and El Roble projects are located in the Cordillera Occidental of Colombia. The Equis vein deposits have sulfide ore reserves of 95,955 mt (105,770 st) grading 8.51 g/mt (0.248 oz/st) gold and 12.4 g/mt (0.36 oz/st) silver with 0.94 % copper, 1.10% lead and 11.06% zinc. Ore reserves at El Roble are estimated as 1.1 Mmt (1.21 million st) grading 4.89 % copper and 3.39 g/mt (0.099 oz/st) gold. Geologically possible reserves at Equis may total 1.6 Mmt (1.76 million st). El Roble has the potential to contain several tens of millions of metric tons. Exploration included diamond drilling, trenching, channel sampling, rock and soil geochemistry, geophysics and tunneling. Erratic high grade samples affect the weighting of channel and drill core samples. Secondary reconcentration of gold is also a factor. Sample maps and graphs show the distribution of values to be highly variable. Geostatistical treatment of some of these data gave important guidelines for parameters such as sampling densities, drill hole spacing and confidence limits.

Jan 1, 1992

By B. Pollack, L. G. Stevens

INTRODUCTION In other chapters of this volume many of the important technical aspects in the design of concentration and dewatering circuits have been discussed. This chapter will deal with the steps necessary to bring that information together in order to produce the specifications necessary to purchase the equipment and complete the design. It deals with the plant as a whole rather than just concentration and dewatering, since the steps are essentially the same. Prerequisites to Engineering Design Feasibility studies will have been completed and design parameters established prior to the start of the engineering design stages for a metallurgical plant. The data base for the feasibility studies will have included laboratory and/or pilot plant test work, historical data, proposed plant site data such as climatology, topographic mapping, geological and geotechnical information. These studies will have established the general project design criteria as well as the process design criteria leading to the preparation of preliminary flowsheet ( s) and/or alternative process flowsheets. These will have incorporated metallurgical and material balances, allowing identification of major equipment items, and development of preliminary plant site plot plans. Preliminary technical and economic assessments and recommendations will have been prepared during this phase for Owner's review and evaluation of the technical, physical, and cost effectiveness of the overall project and its component process, process facilities, and supporting facilities elements. After Ownert s approval and decision to proceed, the engineering design can be initiated, using the above information as the basis. Engineering Design Phase The engineering design of metallurgical plants and their elements are generally developed in two sequential stages, namely: (1) Basic ("Front end") Engineering Design (2) Detail Engineering Design Figure 1 relates these engineering stages to the entire program from feasibility through construction.

Jan 1, 1986

By John C. Nichols

The Ouachita, Ozark-St. Francis, and Mark Twain National Forests are located within the Ozark and the Ouachita Mountains regions of southern Missouri, central and northern Arkansas, and eastern Oklahoma- an area generally referred to as the Ozark-Ouachita or Interior Highlands. These national forests share many similarities in geography, biotic resources, and areas of shared public, market, and user interests. The three national forests have joined together to evaluate the natural resources within the Highlands and to study the impacts of the forests on local and regional economies and on the people who use these national forests and their resources. Other federal, state, and local government agencies and universities have joined with the Forest Service in this project, broadening the assessment area beyond the national forest boundaries to include 107 counties in Arkansas, Missouri, and Oklahoma. The diverse, resource-related data for the as- sessment area address social, economic, terrestrial, and aquatic issues and conditions throughout the Highlands. The assessment report currently is being developed by the Forest Service and will include an accoMpanying CD-ROM. The Ozark-Ouachita Highlands assessment has provided a forum for minerals- and geology-related resources, industries, social significance, and potential to be profiled and discussed along with other critical assessment resources. The state geological agencies in Arkansas, Missouri, and Oklahoma were active contributors to this effort. Their reports and other key data bases developed and used for the minerals sections of the published assessment report will be available either as a supplemental CD-ROM or on the Internet. Updated information on the progress of the assessment report can be found at the Ouachita National Forest Website, at <http://www.fs.fed.us/oonf/ouachita.htm>. This will also include a matrix displaying the minerals in the assessment area, their occurrence by assessment-area counties, and a potential ranking, as well as information on interesting geologic features. The state and federal reports, resources, and data bases, have helped answer key questions concerning mineral resources and related issues in the Highlands:

Jan 1, 1999

By Julius Rubinstein, Victor Badenicov

The development of column flotation is the most significant achievement in the area of mineral processing in the last few decades. The high efficiency of column flotation allows a reduction in process time, a decrease in the number of cleaner stages and in the volume of the circulating load in the flotation circuit, and, as a result of this, an increase in consistency and reliability of operations. Unique multisectional column flotation apparatuses were developed. Such designs have a number of essential advantages as compared to monocells. A change from monocells to multisectional column allows to provide favourable conditions for separation of particles of various floatability. To attain that aim particular hydrodynamics and aeration regime is established in each section of the column, operational variables being adjusted according to the floatability of the material. Therefore in multisectional column machines the material-apparatus feedback can be established, process variables optimized depending on the characteristics of the floated materials. This concept forms the basis of column designs we have developed. For the first time column cell with alternating slurry Rows (combination of co- and counter-current flows) were developed and commercialized . Successful operation was reported of the new generation of columns with cell volumes of 10 to 80 m3 and the height of 4-6 m in copper, molybdenum, antimony, tungsten and nickel ores and coal slurry processing at different plants. Operation experience of these cells showed their considerable design and operational advantages and verified the proposed options and relationships. The new apparatus is provided with a pneumohydraulic aerator allowing to control bubble size distribution, the service life of the aerator is not less than two years. A multilevel flotation model was developed and used to work out a technique of column design and process parameters calculation. A method of the apparatus design parameters calculation based on laboratory test results (scaling up) was worked out. The method is based on flotation macro and micro-hydrodynamics simulation.

Jan 1, 1995

By Nora M. Hutt, Kevin A. Morin

Geochemical static tests are a basic component of geochemical assessments and predictions for acidic drainage, and include acid-base accounting and total-metal contents. A key database for such information is the International Static Database (ISD), which currently contains more than 45,000 static-test analyses from over 200 minesites. Such large databases are useful in identifying general trends and, just as important, the percentage of exceptions. General trends and significant exceptions can be seen in the ISD by comparing one parameter against another. This paper shows that bulk neutralization potential (bulk NP) obtained from acid titration is often a good indicator of neutralizing capacity above values of 10 t CaCO3/1000 t. Other types of calculated neutralizing capacity are then compared to bulk NP to determine if they, in fact, reliably represent actual neutralizing capacity and if they can substitute for the more time-consuming and expensive bulk NP. For example, bulk NP correlates better with Carbonate NP (from carbonate analyses) and Total-carbon NP (from total-carbon analyses) than with (Ca+Mg) CaNP (from solid-phase analyses of calcium and magnesium). This means that, in general, (Ca+Mg) CaNP is a less accurate substitute for bulk NP than the other types, and thus requires intensive site-specific justification if a substitution is attempted. This also means that measurement of the frequently less-expensive parameters of total carbon and carbonate can often substitute, within site-specific error bars, for the bulk-NP titration. However, the ISD also shows that non-carbonate minerals make the primary contribution to neutralization in up to 20% of samples, so Carbonate CaNP and Total-carbon CaNP do not always reveal the full neutralizing capacity of a sample.

Jan 1, 2000

By John Reilly

"INTRODUCTION Successful management of megaprojects, meaning delivery within budget, on time and according to expectations, is a fundamental requirement. Many projects have been successful, some have not. This paper reviews megaprojects1, drawn from the author’s projects over the last 50 years, with “lessons-learned” and a summary of recommendations that, if implemented, would improve megaproject management and delivery. Characteristics of these projects are summarized – including outcomes related to key factors such as politics, management, organizational structure, contracting methods, cost estimates, risk and other elements. The projects include the Washington D.C. Metro, Boston Southwest Corridor, LA Metro, Toronto Rapid Transit Expansion Program, BART San Francisco, MARTA Atlanta, London Underground and the Washington State Highway megaprojects – including the SR520 Floating Bridge and the Alaskan Way tunnel. Other projects that illustrate successful, innovative management and contracting methods include the Madrid Metro (low-cost delivery), the Sydney Northside Storage tunnel (outturn cost, alliancing) and the Lake Mead tunnel (working in partnership, risk management). KEY ISSUES A characteristic of megaprojects is that their complexity increases exponentially with size (Reilly 2010, Galloway et al. 2012). These projects are very visible politically, have many stakeholders who must be satisfied and generally span many years and political election cycles. To be successful they must meet expectations, function efficiently, and be delivered under, at or close to budget and schedule. There is a widely-based perception that large, complex projects are always delivered late, over budget and with problems. Specific examples include projects such as Tren Urbano in Puerto Rico – $1.28 billion over initial budget (+133%), the Silver Line Transitway Project in Boston – $286 million over budget (+90%), the London Jubilee Line Transit Project – 2 years late and £1.4 billion (+67%) over budget, the Channel Tunnel Rail Project – £3.7 billion (+80%) over budget, Denmark’s Great Belt Link rail and road link (54% over budget), the 2003 Woodrow Wilson bridge tender in Virginia (72% over estimate) and Boston’s Central Artery Project – several billion (perhaps 100%) over what would have been a realistic initial budget and years late (Reilly 2001c, 2004a; Salvucci 2003)."

Jan 1, 2016

By Adrian Smith

14.1 INTRODUCTION From the time the last recoverable gold is extracted from ore on a heap leach, the ore - now spent ore - changes from being the resource, indeed the raison d&apos;etre of the operator, to a potential liability. The spent ore contains process chemicals and constituents leached from the ore both in the interstitial pore water in the heap and on the surfaces of spent ore. If these chemical species were released to the environment, it could have a potential impact on the quality of both surface water and groundwater resources. Accordingly, the spent ore must be left in a condition where it does not present a potential environmental liability. This can be accomplished either on a decommissioned pad or, if the heap leach is operated as a load/unload scheme, in a waste facility. In general, the residual cyanide level in the spent ore is the perceived controlling factor when determining whether the leached material is suitable for abandonment. Regulatory authorities set permit conditions for heap leach abandonment in terms of residual cyanide levels, for example, Smith and Struhsacker (1987). This chapter discusses cyanide degradation and methods for detoxifying a heap leach in order to meet regulatory limits/permit conditions for abandonment. Cyanide degradation and detoxification by "natural" processes are described initially, based on the chemistry and geochemistry discussed in Chapter 13. This leads to a review of chemical treatment processes for detoxification of spent ore. The advantages and apparent disadvantages are then appraised. 14.2 NATURAL DEGRADATION AND DETOXIFICATION For the purposes of discussion, "natural" detoxification processes can be divided into two alternatives: • Passive abandonment (the so-called "do nothing" option); and • Water In the first alternative, the spent ore is merely left on the leach pad to age; the second alternative involves the flushing of the spent ore with water.

Jan 1, 1988

By J. P. Mogan

As diesel exhaust contains a mixture of several materials of varying degrees of toxicity which interact to a currently undefined extent in their effects on health, the prescription of a quantity of ventilating air to dilute the exhaust to a safe level is necessarily, to some extent, arbitrary. Some jurisdictions rely on the application of a safety factor to a single "governing" component (usually nitric oxide for most modern diesel engines). Others specify that the total of the "adjusted" concentrations (actual concentration divided by maximum permissible concentration) not exceed "one" for related constituents (irritant gases for example). Neither of these approaches, however, takes account of the interaction of the gaseous constituents with diesel soot (although many studies verify such interactions), nor do they provide any economic incentive for reducing pollutants which are not included in the criteria for ventilation. Because of these factors, the Canadian Department of Energy, Mines and Resources awarded a contract for a study of the current literature bearing on the "Health Implications of Exposure of Underground Mine Workers to Diesel Exhaust Emissions" (1). One goal of this study was the formulation of a new ventilation criterion. The contractors, Ian W. French and Associates, studied more than 1500 literature citations which they interpreted in a 350 page report and summarized in a companion 1600 page annotated bibliography. In the recommendations section of the report, they defined a new &apos;health effects index&apos; or ventilation criterion "V" as follows: [ ] The contractors consider that a "V" between 3.0 and 4.0 poses a moderate threat to health which could be alleviated by the use of protective equipment such-as filter helmets. If V exceeds 4.0, an increase in ventilation quantity to reduce the pollutant levels is indicated. In the case of mines with a quartz content in excess of 20% in the ore, the contractor recommends that a term. [Quartz]/ [Threshold Limit Value Quartz] be added to the equation, and [2.0 = ?v? = 30] be the range justifying protective equipment, with added ventilation if "v" exceeds 3. A degree of pragmatism was inherent in the development of the equation, as the contractors recognized that the components of the index, if it is to be useful in the near term, should be within current routine underground environmental assay capability.

Jan 1, 1980

By Alfred Petrick, Roger Dewey

INTRODUCTION Estimation of capital and operating costs for mines and mills has long been a necessary ingredient of feasibility studies, and though many practitioners have successfully used "rules of thumb" and factor estimating methods for many years, a great deal of work is usually required to prepare an initial estimate and/or to update or revise such an estimate. Development of comprehensive mining and milling cost models during the 1970&apos;s by the U.S. Bureau of Mines and by Canadian workers (O&apos;Hara and Mullar) provided the basis for feasibility-level cost estimates, but the numerous equations and the apparent rigor and complexity involved in preparing a project estimate using these models has discouraged many, if not most, potential users. Their acceptance was also hindered by a broad misunderstanding by many engineers in the minerals industries who had little exposure to such procedures, and by a possibly inadequate level of explanation or documentation of the parametric approach to cost estimating. The preliminary nature of most databases, geologic/engineering/ and market, available to the engineer or analyst during the early stages of a mineral project is a universal problem. Basing corporate mega-investment decisions or even individual equipment selection decisions on the usual "SWAG’s has always been a risky practice, but in the current business environment, the "guesstimates" are almost &apos;verboten&apos;, and are certainly becoming antiquated with the advent of both the described models, the ubiquitous microcomputer, and the nearly-ubiquitous hybrid species known as the "computerized engineer". Engineers have almost always sought techniques of cost estimation that can be applied and justified when geologic and engineering information for a mining project is preliminary and incomplete. The answer, in some cases, is the "parametric" approach. Equations, based on actual, past experience are developed to relate capital and operating costs to major physical variables such as the rate of mining or milling, with modifiers for such details as mining method, process characteristics, methods of construction, or types of equipment . Cost Engineering Standards Five major types of fixed capital cost estimates have been defined by the American Association of Cost Engineers : 1. Order-of-Magnitude estimate based on previous cost data and minimal knowledge. usually a "ratio estimate" with no better than +/- 30% accuracy. 2. Factored or ratio estimate based on major equipment costs with probable accuracy in the 25 to 30% range. 3. Budget authorization or Preliminary estimate, based on sufficient project data to

Jan 1, 1987

By Robert S. Ondrey, Robert A. Haney, Kenneth G. Fields

As part of an MSHA program to assess the effectiveness of dust control parameters for longwall mining operations, respirable dust studies were conducted on six longwall sections. The purpose of these studies was to evaluate the influence of airflow and production on longwall dust control. The sections were sampled for 3 to 35 shifts. In addition to dust, ventilation and production measurements, other dust control and operational parameters were measured. The water spray system at each operation conformed with generally recommended practices. Therefore, improvements to the dust control system would be most appropriately achieved through improvements to ventilation along the face. Two methods of assessing the influence of ventilation and production were used. For two of the six mines where extensive sampling (22 and 35 shifts) was conducted, multiple linear regressions were used to relate the dust concentration versus airflow and production data. For these two mines, statistically significant correlations existed between concentration and both production and airflow. The multiple regressions fit to the data indicated that for every 0.47 m3/s (1,000 cfm) increase in airflow, there would be approximately 0.04 to 0.07 mg/m3 reduction in dust concentration; and, for every 1,000 ton increase in production, there would be approximately 0.15 to 0.50 mg/m3 increase in dust concentration. For the four mines with limited data (3 to 10 shifts), as well as for the two mines with extensive sampling, an average ratio of airflow per ton of coal mined was calculated by proportioning the airflow to production and the designated occupation personal respirable dust exposure to a 2.0 mg/m3 concentration. The airflow-to-tons ratio required to maintain a 2.0 mg/m3 dust concentration ranged from 0.0038 0.0071 m3/s per ton (8 to 15 cfm/ton). Both of these data treatments were used to calculate airflow requirements based on normal production rates. For the two mines where both the regression and airflow-to-tons ratio were calculated, results from the two methods agreed within 25 percent.

Jan 1, 1993

By Alena Conrads, Markus König, Markus Scheffer, Markus Thewes, Hannah Mattern

"INTRODUCTION In recent years, the development of tunnel boring machines has revolutionized the tunneling industry. The general structure of a tunnel boring machine (TBM) can be described as a multi-component system. To achieve high performance rates, machine elements have to perform reliably. Performance losses and interruptions of the construction process should be prevented or at least, reduced to a minimum. In order to achieve high advancing rates, cutting tools must be in good condition at any time. Furthermore, failure of individual tools might result in increased and faster wear of remaining structural components. For this reason, maintaining and replacing cutting tools is crucial to prevent project delays due to insufficient boring performance. Precise prognoses concerning the prevailing geology are infeasible and thus, efficient scheduling of maintenance actions is challenging. The intention of this work is to analyze different approaches of TBM maintenance in soft ground. Different strategies are analyzed by the use of a simulation approach. Cutting tool condition is regarded as performance limiting factor. The maximum operation time of the cutting tools is determined using the “Soil Abrasivity Index”. A case study serves to show the effects of different maintenance strategies on total project duration. SCHEDULING OF MAINTENANCE ACTIONS Scheduling maintenance actions of single and multi-component systems concerns numerous fields of industrial production. Concentrating on mechanized tunneling in soft ground, cutting tool maintenance represents the most common reason for intermission of advance (Köppl 2014). In literature, time-based and condition-based maintaining techniques are widely discussed. The following section outlines the principles of time- and condition-based policies. Furthermore, examples for their applications in recent research are given. Time-Based Maintenance In general, the time-based maintenance approach assumes that the failure behavior of the system components is predictable. Maintenance actions are scheduled according to predefined intervals (e.g., hours, days, weeks). A detailed analysis of preventive repair times or interval forms the basis for maintenance decisions. For this approach, an adequate quantity and quality of data is fundamental. Numerous studies deal with the analysis of time-based maintenance for production equipment. In fields of mechanized tunneling, unplanned machine downtimes restrict the applicability of time-based maintenance approaches. Thus, this paper mainly focuses on condition-based maintenance."

Jan 1, 2016

By R. Bruce Thorndycraft

Heap leaching should not be chosen for development of a property solely on the basis of paper studies. Management should be made to understand where the risks are in heap leaching, and management should be alerted to the variability of costs in heap leaching, many of which are based upon company philosophy. These concepts are complex, but easily illustrated with guided tours of existing heap leach facilities.

Jan 1, 1998

By James F. Archibald, Ferri Hassani, Jacques Nantel, David O. DeGagné

Operators depend upon the success of backfilling programs to ensure that mine activities run continuously (Archibald et al, 1993) where backfilling is an integral part of the mining cycle. In order to ensure that the backfill is prepared and delivered to the stope possessing the required engineering properties, as designed, and that the backfilling process is safe, mine operators at all stages of the backfilling system must be adequately trained for their tasks and be aware of how their work affects other backfilling stages, both up- and down-stream of their station. Mine operator training is primarily based on learning on-the-job from experienced supervisors or training personnel. Classroom training and the availability of reference manuals are less common. While in use by many industries for some time now, computer-based training (CBT) is becoming more readily accepted by the mining industry for a wide variety of training. It is not intended to replace on-the-job training, but to augment it. CBT offers many advantages to enhance operator learning by utilizing multimedia presentations, interactions and on-line testing to teach complex concepts and relationships. Discussions with many operations revealed that there is an industry need for a backfill guide for frontline operator training and as a reference manual for operators and mine engineers. Sponsored by the mining industry and the CIM, "Mine Backfill, An Operator&apos;s Guide" has been developed by the Departments of Mining Engineering at Queen&apos;s University and McGill University to fulfil this need. "Mine Backfill, An Operator&apos;s Guide" is an interactive multimedia package that incorporates high quality colour images, animations, sound, and video with text to review the main backfill methods employed in Canada, including hydraulic/slurry, paste and rock fill. The Guide is divided into training and reference primer sections. When in training mode, the operator is guided through a presentation of common mining and backfill practices and technologies so that he/she may become more familiar with different aspects of backfill processes. To ensure that key concepts have been learned during training, the operator will be tested after each section using a series of multiple choice and true/false questions. When in reference mode, alternatively, the operator may browse through the information in a more detailed and independent fashion through a series of interactive menus. The Guide focuses on the following main topics: resources, surface preparation, distribution, stope activities, health and safety, and quality control. Special features included in the Operator&apos;s Guide are a Buyer&apos;s Guide, a Trouble Shooting Guide, and an interactive glossary of backfill and mining terms. Concepts in interactive and computer-based learning are also reviewed in this paper.

Jan 1, 2001

By C. D. Elifrits, N. B. Aughenbaugh

Introduction Federal and state laws enacted to regulate coal mining and the accompanying public concern about the adverse effects that mining might have on land use have focused much attention on the subsidence of the land surfaces over underground mines. Some questions asked are • Will abandoned mines subside in the future even though they have exhibited no surface adjustment to date? If so, when will subsidence take place? • Can a specific surface area be affected by subsi¬dence more than once? • What surface form will the subsidence feature take? • Is there any way to predict when subsidence will occur and end? An attempt is made here to address these questions insofar as it is possible with available data and experience of evaluation of subcritical extraction conditions. Mining that results in subsidence features can be classified as supercritical, subcritical, and critical with respect to the minimum span lengths of the worked-out rooms (Fig. 1). A complete basin will not develop fully immediately over subcritical conditions because the span lengths of the entries are less than the critical dimension (Wardell, 1983). Passageways, entries, and room-and-pillar panels are all included in the subcritical category. Because of the short span lengths, subcritical mining conditions tend to be stable over the life of the mine unless adverse geologic conditions exist (Gray et al., 1974; Dunrud, 1976). Various roof control and support procedures are employed to maintain entry stability during mining. Little information exists concerning the time factor of subsidence over areas of subcritical extraction. Data that have been collected relate to longwall mines, whereby the cavity sizes are designed large enough to induce caving behind the working faces so that complete subsidence at the surface occurs within a short period of time. The UK&apos;s National Coal Board (1975) has reduced the enormous amount of data it has collected to empirical correlations and has published these correlations in its Subsidence Engineers Handbook. The handbook states that "None of the foregoing remarks applies to pillar and stall working," with respect to the section discussing the time factor of subsidence caused by longwall mining (National Coal Board, 1975, p. 40). To understand and evaluate how time relates to subsidence above room-and-pillar mines, one must first recognize how mine cavities deteriorate and fail. The propagation of failure to the surface must then be studied, and, finally, the failures that are manifest at the surface in the form of depressions accompanied by both compressional and tensional movements must be investigated. Failure of mine cavities Any cavity beneath the earth&apos;s surface represents a condition of nonequilibrium and tends to close with time. Failure of cavities in a subcritical mine can be classified as failure of the roof, failure of the floor, failure of the supports (pillars, cribs), and elastic rebound of the rock. One mode of failure may dominate, but generally combinations of the four will occur. Also, initial instability of geologic materials and structural features and in situ stress may be unknown factors that can contribute to cavity failure. Roof failure is the most obvious opening stability problem in most operating mines. The bridging capabilities of the roof rock are dependent on the rock&apos;s material strength, geologic discontinuities, and reaction to humidity changes in the air of the mine. This type of failure may vary in magnitude from gradual flaking to massive roof falls. Mine floor instability can be manifest in the form of a general heave over a broad area or of individual pillar punching. The weaker and thicker the immediate floor material, the more prone a mine is to experience floor problems. Water seeping into an area can soften the underclay, reduce its bearing capacity, and precipitate failure of support for the overburden.

Jan 1, 1987

By M. M. Botz, F. W. DeVries

Discussion by F.W. DeVries Comments on natural cyanide attenuation. The paper by Botz and Mudder in the November 2000 issue of Minerals and Metallurgical Processing had one disappointing aspect. Table 1 (p. 229) failed to include a significant species - neutral copper cyanide (CuCN). This compound has been most trouble-some in many plant-scale cyanide cleanup processes, includ- ing acidification-volatilization- reneutralization (AVR), hy- drogen peroxide cleanup and air/SO2 cleanup. It has also plagued many pilot efforts to use ion-exchange (IX) resin technology. This author elucidated the concept that, under some conditions, copper cyanide failed to precipitate out of treated effluents and appeared to exist as a filterable colloid that reported as copper and cyanide, with the latter measured as WAD (weak-acid-dissociable) cyanide in that analytical method. The concentrations measured were very near the levels of 0.2 ppm, often the regulatory limit for WAD cyanide. The copper concentration, using the 1 : 1 ratio for this com- pound, would then be more than 0.4 ppm, which could also be unacceptable. We discovered that it was unsafe to assume that the Cu(CN),- would surrender both cyanide ions simultaneously and allow cupric copper to be removed subsequently as, say, cupric ferrocyanide (Hatchett&apos;s Brown). This compound could be ignored as safely as the "iron blues," as essentially an inert solid, stable in geologic time if not exposed to sunlight. Under poorly controlled oxidation of the cyanocuprite ions (copper with four, three or two cyanides complexed to the copper), CuCN could well be formed and could not be assumed to precipitate out as a solid. The authors of the paper, by failing to include the CuCN species in their model, may have left users of the model prone to ignore the possible occurrence of the colloidal(?) form in the discharge, thus finding an effluent that failed to meet regulatory limits. (See addendum below.) This somewhat ephemeral material may well have been responsible for some large-scale failures of cyanide destruction technology to succeed at sites in North America and elsewhere. Two specific references alluding to this occurrence may be found: Griffiths et al., 1987, "The detoxification of gold-mill tailings with hydrogen peroxide," Journal of the South African Institute of Mining and Metallurgy, September, p. 282. DeVries, 1989, "Improved understanding in detoxification of cyanide leach liquors using peroxide" Rand01 Gold Conference, Sacramento CA, p. 15 1. It is advisable for workers in this area of cyanide removal or recycle to avoid assuming that CuCN will, with certainty, precipitate out of solution, or that the dicyanocuprite ion will part simultaneously with both cyanide ions and allow cupric copper to be dealt with by other reactions. One further piece of work may be incomplete. That is, a definitive study of the long-term behavior of CuCN that does precipitate when exposed to other compounds present in neutralized tailings. Addendum. The extremely complex reaction of cupric ion on cyanide- which, of course, results in oxidation of cyanide to cyanate, probably through the intermediate stage of forming cyanogen [(CN)2] from oxidation of HCN with its hydrolysis to cyanate and cyanide ion - involves two cyanide moieties and the reduction of the cupric ion back again to the cuprous state. This, then, can form the CuCN potential "solid" that was already the undesirable species. In cyanide cleanup, with either air/SO2 or hydrogen peroxide, copper is generally employed as the catalyst, so it is unclear what happens when the CU(CN)2-1 is ultimately exhausted. Certainly, the field evidence is quite clear that all

Jan 1, 2002

By J. E. Wennen

Screens serve several important functions in CIL and CIP processes for gold extraction. One important screen application known as interstage screening allows slurry to be continuously transfered through a series of CIL or CIP tanks while retaining activated carbon in the tank. In some cases, this has been accomplished by using airlifts to carry slurry containing activated carbon to an external screen which separates carbon as oversize and returns it to the tank by gravity. The undersize slurry advances to the subsequent tank. More recently this has been accomplished using launder screens and Kambalda type screens. These are essentially static screens that rely on air blasts or mechanical sweepers to clear carbon away from the screen surfaces.

Jan 1, 1998

By Fabian Laumen, Sunny Schoone, Felix Leaman, ELISABETH CLAUSEN

The paper evaluates the attenuation behavior of acoustic emission (AE) signals in four different materials used in marine environments. The study aims to identify materials with the highest and lowest AE signal attenuation, providing insights into wave propagation and structural monitoring in deep-sea mining applications.

Dec 1, 2023

By G. T. Newell, J. G. Stone, V. M. Mejia

Introduction Advances in drilling have reached a point where large diameter cores can be recovered from "tight," or weakly indurated placer gravels. In such ground, core drilling can provide more reliable data regarding tenor than can be obtained using churn drilling or similar classical techniques. It can also provide metallurgical and geological information that is not available from samples obtained through alternate methods. In 1985, Coastal Mining Co, a subsidiary of M. A. Hanna, and Western Gold Reserves began to review a Tertiary placer deposit owned by San Juan Gold at North Columbia, CA, about 14 km (9 miles) northeast of Grass Valley. The deposit is one of the largest remaining unmined portions of the formerly extensive early Tertiary ancestral Yuba river system. It has been known since the 1850s, has been the subject of much technical literature, and has been the object of at least four previous drilling programs. The eastern one-third of the 6 km (3.7 mile) stretch of the channel between North Columbia and Badger Hill was partially stripped by large scale hydraulic mining in the late 1870s and early 1880s. Mining ceased in 1884 when the Sawyer Decision prohibited further discharge of hydraulic tailings into the Sacramento and San Joaquin Rivers. By that time, about 30 to 45 m (100 to 150 ft) of relatively low grade upper gravels had been removed over some 81 hm2 (200 acres). About 90 to 105 m (300 to 350 ft) of higher grade middle and lower gravels were left at least partially stripped. In 1914, a few churn holes were drilled along a widely-spaced line. In 1938-1939, Selection Trust conducted an extensive drilling campaign to evaluate the deposit. Particular attention was directed toward the partially stripped eastern portion. In 1968, the US Geological Survey drilled three churn holes in the eastern part of the deposit. The US Bureau of Mines conducted experimental mining and drilling in the Badger Hill area. In the late 1970s, Placer Service Corp. acquired a lease on the deposit. Between 1979 and 1984, Placer Service drilled 28 large diameter BADE (a German-manufactured machine) drill holes on the eastern portion of the deposit. The surviving records from the widely-spaced 1914 drilling program are fragmentary and the reported grade not well substantiated. The 1968 holes were drilled for scientific purposes. Again, drilling details are not available. However, detailed records for both the churn drilling program and the BADE program were available and formed the basis for the initial evaluation of the property. Geology The geology of the auriferous Tertiary gravels of California have been described by Whitney (1880), Lingren (1911), and, more recently, Yeend (1974). In general, the Tertiary gravels in the North Columbia area occupy a broad channel cut into pre-Tertiary igneous and metamorphic rocks. The upper, or white gravel is overlain conformably by volcanic tuffs and volcaniclastic rocks. A middle gravel is characterized by the presence of silicified and carbonized wood. A lower blue gravel unit has relatively coarser cobbles and contains a higher proportion of igneous and metamorphic cobbles than the other units. The upper gravel consists of interbedded pebbly sand and silty, or clayey sands with prominent cross bedding. Most of the pebbles are well rounded and consist mostly of white vein quartz and quartzite. The upper unit is moderately well compacted. Exposures in the walls of the old hydraulic mine pits stand at 45° and 50° angles. The gold content of the unit is well below an economic cutoff. The middle gravel - included with the upper unit by Yeend (1974) - is coarser grained, with carbonized wood, and 75 to 100 mm (3 to 4 in.) cobbles of metased-imentary and metavolcanic rocks in a sandy matrix containing abundant lithic fragments. The upper contact appears to be conformable, but the lower portion of the unit appears in places to consist of reworked lower gravels. The unit contains less clay than the upper unit and is somewhat more friable than the underlying lower gravels. The gold content, while somewhat higher than the upper level, is too low to be of ore grade. The lower gravel averages between 30 to 45 m (100 to 150 ft)

Jan 9, 1988