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By James A. Finch

The Gaudin lecture is devoted to fundamentals. Given my involvement with column flotation over the last 20 years, this is my topic. I have tried to identify spinoff basic research and provide a personal list of favorites. Two are selected for discussion here. They are single bubble rise and bubble surface area flux. The rationale for such work, as well as some of the findings, are reviewed. Support for fundamental work seems as volatile as the economy (and the link is probably not coincidental). I hope some utility to the minerals industry is evident in the text. Column flotation in its current form originated in the early 1960s. The method was invented by Boutin, patented by Boutin and Tremblay and pursued doggedly by Wheeler. Their aim was to overcome the problem of mechanical entrainment of fine particles in the water carried by the bubbles, which they were convinced contributed to lowering flotation concentrate grades (Wheeler, 1988; Finch and Dobby, 1990). The solution was a combination of columnar geometry, the addition of wash water into a deep froth and the maintenance of an excess of water to the tailings, as compared to the feed (a so-called positive bias). Success at Les Mines Gaspe, Quebec, on cleaning Mo in the early 1980s signaled the start of industrial acceptance, making the column one of the flotation machines of choice today. Over the next 20 years, column applications mushroomed to cover virtually all mineral commodities, as well as nonmineral applications such as soil decontamination, deoiling of water and deinking of recycled paper. In the wake, research groups are continuing to probe the fundamentals, conferences and workshops have been transferring the technology and specialist companies have emerged to service the demand. Innovations were triggered as problems arose (for instance, in gas injection systems) and novel flotation cells, much in vogue in the 1990s, had the door opened. Mechanical cell manufacturers have not been idle - there is nothing like competition to provoke a fresh look at an established business. Flotation-cell technology has probably not seen so much activity since the patent rush in the early days of flotation, which occurred around the turn of the century. For an industry often considered reluctant to risk innovation, this is not a bad record. A good idea will find a place, but patience (not always an innovators strong suit) and careful preparation are required. Two books (Finch and Dobby, 1990; Rubenstein, 1996), three symposia (Sastry, 1988; Agar and Huls, 1991; Gomez and Finch, 1996), and a series of reviews (Dobby and Finch, 1991; Finch and Dobby, 1991; Finch, 1995; Finch et al., 1995) document the period. Like many good ideas, column flotation has been overexposed, and there have been disappointments. In a recent example, columns performing scavenging duty were replaced by mechanical cells (Chong et al., 1998). The further applications stray from the original - to

Jan 1, 1999

By John B. Jones

New concepts in shaft kiln processing are included in the design of two large diameter shaft kiln plants now under construction. One kiln will be used to produce cement clinker at the Boettcher plant of the Ideal Cement Company at Fort Collins, Colorado. The other kiln is for the semi-commercial oil shale plant for Petrobras, the Brazil National Oil Company at Sao Mateus do Sul, Parana, Brazil. Investigations on application of the new concepts are proceeding in a portable pilot kiln. The shaft kiln has long been recognized as a method of gas-solids contacting having the following principle advantages: l. High thermal efficiency 2. Ability to handle a variety of feed materials with a wide range of particle sizes 3. Minimum investment and operating cost

Jan 1, 1963

By P. L. Dunn

Automatic control of fine coal cleaning circuits has traditionally been limited by a lack of sensors. Although several nuclear-based analyzers are available, none have seen widespread acceptance. In this investigation, image analysis is used for on-line determination of ash content in fine coal slurries. The ash content of a flotation tailings stream is correlated to the mean gray level of the slurry. Based on a 90% prediction interval, an analysis of sensor performance indicates an accuracy of 73 ± 4% ash over a range from 65-90% ash. While this result is comparable to existing nuclear-based analyzers, the video-based system is only a fraction of the cost. The sensor is currently installed at the Middle Fork preparation plant owned by Pittston Coal Company.

Jan 1, 1997

By A. D. Davis

The disposal of spent ore from cyanide heap-leach processing facilities is of concern to the mining industry, regulatory agencies, and the public. Disposal of several hundred million tons of additional spent ore is planned in the western United States during the next decade, with increased amounts likely if gold prices rise. In the past decade, more than fifteen million tons of spent ore have been disposed of in the Black Hills. South Dakota's off-loading criterion for cyanide has been set by the state's Department of Environment and Natural Resources at 0.5 parts per million (ppm) as weak acid dissociable (WAD) cyanide. Nitrate off-loading limits vary for each mine in the state, depending on ore type and other factors. Nitrate concentrations in leachate from spent ore depositories could be caused by conversion from other nitrogen-containing species. One likely source of nitrate is residual explosives from blasting of ore. Previous work on the combustion products of nitrogen explosives has shown incomplete 3xidation (Johansson and Persson, 1970). Another likely nitrate source is degradation of cyanide to ammonia, followed by oxidation to nitrite and nitrate.

Jan 1, 1994

By A. Adhikari, P. Tukkaraja, S. Jayaraman Sridharan

Shafts are critical components of the mine ventilation systems and contribute significantly to the mine ventilation pressure. Estimation of shaft pressure losses is an important aspect of mine ventilation planning and shaft frictional pressure losses are extensively studies. Shock losses contributed by shaft buntons are estimated using the interference factor model developed half a century ago. This study aims to develop a better performing prediction model for shaft buntons using state of the art experiment and numerical data, and machine learning techniques. Results show that the present prediction model is performing poorly, and the prediction performance can be improved by incorporating additional predictor variables. The Artificial Neural Network standardized Machine Learning model developed in this study is able to predict the shaft bunton shock loss with high accuracy (R2 = 0.956).

Jun 25, 2023

By J. B. Ackerman, C. S. Bucans

The Sunshine Mining Co. has recently started up its new silver refinery in Kellogg, ID. The new plant recovers silver anc copper values from a high grade tetrahedrite concentrate using hydro- metallurgical processes. The plant and process are both new and have gone through the initial startup phase during the first six months of 1985. The full scale plant was developed from laboratory test work. While the metallurgy of the process worked as expected, the material handling problems were difficult to over-come. As might be expected, those parts of the plant that represented available technology started up easily, while the parts that represented new technology presented the most difficulty. Topics covered in this paper include: 1) insufficient type and amount of agitation in the leach vessels, 2) discharge of a non- homogeneous slurry from the leach vessels, and 3) problems associated with very low flow rates in slurry lines.

Jan 1, 1987

By David S. Bolin, Bruce Kennedy, Leigh A. Readdy, Graham Mathieson

INTRODUCTION The estimation of ore reserves is a process that begins with the earliest exploration stages on a property and continues throughout any subsequent evaluation and exploitation of the deposit. During exploration and preliminary evaluation, the results of these reserve estimates constitute the basic data for prefeasibility studies and economic analysis. The decision to continue exploration and development or to abandon a prospect is often based upon these studies. During the active life of a mine, reserve computations are continuously revised to assist in development planning, cost and efficiency analyses, quality control, and improvement of extraction and processing methods. Accurate reserve estimates are also required when financing a project, purchasing or selling a property, and for accounting purposes such as depletion and tax calculation. It is important to remember that the reliability of ore reserve estimates varies progressively through time as more and more information becomes available. The lowest order of reliability of estimation of reserves exists at the time of discovery. The maximum level of certainty concerning the ore reserves within a deposit is reached when the deposit is completely mined out. Between these two extremes are variable levels of certainty as to the tonnage and grade of the resource. This is particularly true of that portion of the mineral resource that constitutes the "minable reserve," as this portion is dependent upon economic as well as geological and technological factors. In the following discussion, several of the factors affecting ore reserve computation and some of the commonly used methods of calculation are presented. The first part of the discussion is confined to classical methods of hand calculation utilizing level maps and sections. The second part of the discussion presents an overview of somewhat more sophisticated methods of geostatistics which have been developed within recent Years. Geostatistical methods frequently provide the best ore re- serve estimates for many deposits. However, a digital computer is required for such geostatistically derived estimates due to the number and complexity of the cornputations involved. For many properties the classical nongeostatistical methods are adequate, particularly in the early exploration stages. In fact, it is the author's contention that even at those properties where the geostatistical approach is employed for reserve estimation, the geological and engineering staffs should prepare an estimate by traditional means. This provides an internal audit for the reserves and at the same time requires a continuing close appraisal of geologic problems that influence the presently known geostatistically estimated reserves and production at the mine.

Jan 1, 1982

By Andrzej Frycz

The development of the theory of reliability creates more and more possibilities of its application in different domains of technics. The methods evolved to date, both theoretical analysis of the engineering systems and investigations techniques in practice can be put in use a d translated from one science to another one [1,3,7J . It needs to consider the specific of the different sciences. But it is necessary to take account of the specific characteristic of each line of science. The reliability investigations needs, justified in technical and economical viewpoint appear too in mine ventilation technics and systems. The mine ventilation system is a complex of workings and ventilation plants, which serve to supply the networks with a sufficient quantity of fresh air from the atmosphere and then carry off the vitiated air. The condition necessary for a ventilation system to perform definite task is the correctness of all its elements, that is a capability to preserve in a expected range of tolerance, the values of basic parameters. The element of a system which ceases to work correctly is called failed. The basic parameter of all elements of a ventilation system is the characteristic curve, pressure loss - air quantity relation ?p = f(Q) (1) This curve concerning the workings and ventilation stoppings is given by the formula: ?p = R Q2 (2) thus the resistance R (of the workings or stoppings) may be taken as a characteristic parameter. The course of this characteristic f(Q) will be given in a complex form in concern with fans, ejectors, air curtains. The failure of a ventilation plant or working is therefore an event, relying on a change of its characteristic, thus for a definite time the losses of the plant's or working's capability set in for the fulfilment of its function in the ventilation system. The failures are divided in viewpoint of their nature on sudden and progressive. We are in concern with the sudden failures which are characterised by a violent change of the ventilation element parameter for example a fall of roof, or silt up of the working, opening or failure of the ventilation stopping, the stand-off of the fan or ejector etc. We distinguish two kinds of failures on the viewpoint of their action on the air flow: the "opening" failure and "closure" failure. The prevalence of factors, which causes the failures is in connection with the intersection in the mine of roads of haulage, roads of underground transport of materials and lining elements, roads of people motion with airways, as also depends on the rock pressure reaction on the workings lining and plants placed there. Other factors will yet be decisive upon supply decay of fans, ejectors and curtains. The appearance of failure of the ventilation element is therefore of random nature and determines the change of the element, when it is passing from operating state to unoperate. Qualifying the ventilation system element to one of these states is easy, when the characteristic change is clear and occurs suddenly, that was assumed. The investigations conducted by Patrushev and Karnaukh [91 in concern with the ventilation stoppings resistance change in the case of their opening, had shown the possibility of resistance value change in the range of several hundreds higher.

Jan 1, 1980

By B. N. Conley, K. D. Luxbacher

"The underground coal mining industry in the United States has recently seen the occurrences of several high profile, multi-fatality events. The explosions that occurred at the Sago Mine in 2006, the Darby Mine in 2006, and the Upper Big Branch Mine in 2010 have caused a ripple in an otherwise steadily improving safety record. These events transpired in the midst of an unprecedented level of government regulations and modern safety technologies. The recent increase in fatal events in conjunction with a minimal decline of both fatal and non-fatal injuries over the past decade may signify that current safety practices have reached a level of diminishing returns. Risk management, a safety approach that has been successfully applied in various industries including mining across the world, may provide a means to surpass the safety plateau in the U.S. RISKGATE is an Australian risk management program for mines that shows great potential for application in the U.S. However, fundamental differences between the coal mining industries in Australia and in the U.S. prevent direct implementation. This paper discusses aspects of the RISKGATE body of knowledge that require some adaptation before this program may be applied to the U.S. mining industry. INTRODUCTION The underground coal mining industry in the United States has recently seen the occurrences of several high profile, multi-fatality events. These events transpired in the midst of an unprecedented level of government regulations and modern safety technologies. The recent increase in fatal events in conjunction with a minimal decline of both fatal and non-fatal injuries over the past decade may signify that current safety practices have reach a level of diminishing returns. In order to combat this stagnation in performance, a modification to the manner in which the mining industry approaches worker safety may be warranted. The implementation of a risk management based approach presents a potential method for not only surmounting the current safety plateau but also for achieving the long desired objective of zero harm. RISKGATE a risk management body of knowledge designed for the Australian mining industry that shows great potential for application in the U.S. This paper discusses aspects of the RISKGATE body of knowledge that require adaptation before implementation in the U.S. mining industry. These adaptations are based on differences in operating practices and governmental regulations. This discussion will focus on the three major U.S. risk areas of explosion prevention, ground control, and moving equipment."

Jan 1, 2015

By Zhen-xing Diao, Feng Zhao, Ya-dong Xue

"Two important aspects of excavation performance of TBM are advance rate and cutter replacement/ring consumption, which is closed to the cost and construction duration of the project. To achieve this, a database of TBM parameters and the geological-geotechnical conditions during the excavation of Yinhanjiwei was compiled and analyzed. Relationships between different geological and TBM operational parameters were investigated. Results of analyzes show that there are qualitative correlation between Field Penetration Index (FPI) and rock geological parameters. And strong relationships between FPI and operational parameters are suggested. In this study, the cutter replacement and ring consumption were analyzed using statistical method for preliminary estimation in similar conditions. 1. ITRODUCTION Hard rock TBM is widely used in tunneling project nowadays because of its advantages to conventional drilling and blasting, especially in long tunnels. The biggest advantage is its high advance rate. However, rock conditions of tunnels are complicated. So TBM performance becomes a primary topic in tunneling engineering, which includes of rate of penetration, utilization, daily advancement, cutter consumption and so on. Many prediction models have been developed to estimate the TBM performance in the project planning phase. The most widely accepted and used models around the world are semi theoretical CSM model (Rostami and Ozdemir, 1993; Rostami, 1997) and the empirical NTNU model (Blindheim, 1979, Bruland, 1998).With increasing uses of TBMs in different ground conditions, an increasing number of researchers have developed new models in order to predict the performance of TBM more accurately in similar geological conditions.(Gong and Zhao, 2009; Hassanpour et al.,2009, 2010, 2011, 2014; Delisio et al., 2013; Delisio and Zhao, 2014)"

Jan 1, 2016

By M. A. Sochocky

Rod mills were developed, as a logical substitute for the rolls crusher when larger tonnages had to be handled, on a continuous basis. Today, with even larger tonnages to process, the trend is to install larger equipment, automate the plant and develop the highest availability possible, in order to cut costs. The rod mills, because of design limitations in size and speed 1'2, are not able to meet these new requirements. The autogenous mill was a timely development and serves as an excellent alternative for ores that are suitable to this type of grinding. When ores are not amenable to autogenous grinding, selection of a flowsheet is dictated by liberation characteristics and unit operations to be used for mineral recovery.. The following alternatives are available: (a) Multiple rod mills, capable of drawing no more than 2000 HP each, followed by ball mills. (b) Gyradisc crusher followed by two stages of ball mill grinding.3 (c) Tertiary crushing followed by closed circuit single stage ball mills grinding to approximately -100 mesh, as in the copper industry. (d) Tertiary' crushing followed by two-stage, ball milling. Each stage will operate in closed circuit with either screens or cyclones. It is this last alternative, particularly as it

Jan 1, 1972

By Martin Wehrli, Oliver Schneider

"MOTIVATION Tunnel construction companies must keep costs low, finish projects in less time and still meet the high quality and safety standards. Therefore, their main goal is always to optimize the tunnel excavation process. Traditional geodetical methods are used since more than 100 years to ensure an exact breakthrough of a tunnel. Evolving sensors, sensor fusion and increasing computation resources opened new fields of application for geodetic solutions in tunneling. These solutions offer new ways to reduce costs and risks in tunneling. INTRODUCTION Nowadays in tunneling, geodetical measurement methods are not only used for guidance, but also for various other applications. This paper gives an overview about where geodetical measurements are used today and highlights their benefits for safety and costs aspects. Furthermore, advanced solutions are introduced. In progressive projects, they have already become state of the art technology and are in use daily. STANDARD GEODETICAL MEASUREMENT APPLICATIONS In drill & blast tunnel construction processes, there are many standard geodetical applications used in the different construction phases, each having different requirements regarding accuracy. In the excavation construction stage, guidance tasks and setting-out of various positions is done by the foremen. Thanks to simplified measurement tools, the foremen are enabled to carry out these measurement tasks themselves without requiring deep surveying knowledge. In contrast, deformation analysis and network measurement campaigns need sophisticated planning and a profound education in surveying. Therefore, these tasks are carried out by surveyors. Profile control and as-built documentation is required in each construction phase. Depending on the requirements of the project, this task can be carried out either by the foremen or a specialized surveyor. Table 1 provides an overview over typical surveying tasks in an underground project including information about who is doing the job and the according accuracy requirements. The listed tasks are described in more detail below."

Jan 1, 2016

By R. Kingman

At Tahawus, New York, National Lead Company operates the MacIntyre Development, This operation produces 10,600 long tons of ore per day for crushing, 8,000 tons for grinding and processing, of which 2,000 tons per day reports as tailing for disposal in an impounded tailing storage area. With the change in location of mining operations at the mine of National Lead Company, a change in the area for tailing disposal by gravity became necessary, This called for the diversion of tailing from near the proposed mining area to areas more favorable for long term storage by gravity. In order to avail the operation of maximum storage for the future, a uniform minimum grade launder was proposed with a minimum velocity and maximum outfall at the discharge. Disposal of the gravity and flotation tailing was formerly accomplished in open ditch from the concentrator at a grade of 1.5 percent. It was desirable to design the new launder at grades less than this for the greater storage elevation.

Jan 1, 1966

By L. H. Munson, D. E. Hadlock, L. F. Munson, R. L. Gilchrist, P. D. Robinson

All uranium mills are required to perform sampling and analysis for radioactive particulates in their gaseous effluent streams and in the environment. Pacific Northwest Laboratory was requested by the U.S. Nuclear Regulatory Commission (NRC) to provide technical assistance to them for their Uranium Mill Health Physics Appraisal Program. In conducting appraisals, air sampling methods used at NRC-licensed mills were reviewed and several deficiencies noted. This paper includes only environmental and effluent particulate sampling although much of the information is applicable to both in-plant and environmental samples. First, the components of a proper sampling program are discussed: program objectives, program design, sampler design, analyses, quality assurance, and data handling. Then the specific deficiencies, or the "pitfalls" from the first 8 mill appraisals are discussed. The first consideration in establishing an air sampling program is defining the objectives of the program. What is air sampling suppose to accomplish? Many of the deficiencies we have observed have resulted because the desired objectives were not clearly established in the minds of the radiation safety staff. PROGRAM OBJECTIVES An environmental air sampling program ought to fulfill the following seven objectives. The first is to: 1) [demonstrate regulatory compliance]. Although a goal of most programs, regulatory compliance, is not well understood. One has not only to comply with the conditions of the source materials licensee, but one must also demonstrate compliance with 10CFR20 and 40CFR190. For example, 10CFR20.106 states: "A licensee shall not possess, use, or transfer licensed material so as to release to an unrestricted area radioactive material in concentrations which exceed the limits specified in Appendix B, Table II of this part .... For purposes of this section, concentrations may be averaged over a period not greater than one year." Even if a mill's license does not require sampling at the site boundary of maximum concentration, a sample may be necessary to demonstrate compliance with 10CFR20. Most mill personnel are painfully familiar with 40CFRl90.10, which states: "Operations.... shall be conducted in such a manner as to provide reasonable assurance that: (a) The annual dose equivalent does not exceed 25 millirems to the whole body.... of any member of the public as the result of exposures to planned discharges of radioactive materials, radon and its daughters excepted... from uranium fuel cycle operations..." This means a licensee's sampling program must give "reasonable assurance" that the member of the general public receiving in the most exposure gets no more than 25 millirems per year. The sampling program necessary to provide that assurance may or may not be a license requirement. However, merely meeting the license requirements and the explicit regulatory requirements does not necessariarly ensure an adequate effluent and environmental air sampling program. The second objective of the environmental air sampling program, is to 2) [identify the source(s) of contaminants]. This will include not only the routine program, but special sampling for verification of sources and nonsources. Only after sampling can a mill operator be assured that roof vents, laboratory hoods, and other localized ventilation systems are not making a significant contribution to environmental releases. An environmental sampling program should also allow the mill operator to fulfill the third objective, to 3) [estimate exposures]. Even before 40CFR190, a sampling program should have provided the mill operator with the information necessary to determine the dose to the "fence post" person, or at least to determine if doses were well below the 10CFR20 limits previously allowed. The program should 4) [detect and measure unplanned releases]. If there is a fire, a scrubber failure, or if a drum of yellowcake breaks open, measured releases will almost always be lower than conservative estimates. Whether or not a system to provide sampling during accidents is needed is almost always a cost-benefit decision. In general, uranium operations do not sample just in case an accident may occur. Yet they may decide on continuous air sampling in lieu of intermittant sampling partially because of the potential for accidents. Another objective of air sampling is 5) [to provide information on the effectiveness of control systems]. This is always a concern with new or modified equipment and may dictate sampling frequency in other situations as well. For instance, if a small leak in a bag filter cannot be detected by other means, then more frequent stack sampling may be indicated. A routine effluent and environmental monitoring program should also fulfill the sixth objective,

Jan 1, 1981

By M. A. Giralico

Solvent Extraction (SX) technology has undergone significant refinements including outright process changes to achieve greater copper production rates with lower organic entrainment levels. LIGHTNIN has completed a detailed study aimed at improving on the conventional Holmes & Narver and Davy Power Gas type impeller designs1 by optimizing the mixing through the use of high efficiency pumper turbines instead of making system changes as others are doing. Since each individual installation requires a different design head and flow, LIGHTNIN has developed the R320 "family" of high efficiency, curved blade pumper turbines. These impellers are designed to maximize head and flow while reducing entrainment through lower power consumption, less shear generation and turbulence, elimination of air incorporation, production of the optimum droplet distribution, and lower tip speeds. These improvements also maintain the required mass transfer and copper recovery of the traditional impeller designs. The R320 impeller benefits have been quantified using Laser Doppler Velocimetry (LDV) and proven in multiple domestic and foreign installations at flow rates exceeding 15,000 GPM. Design comparisons of this improved impeller type versus traditional technology impellers in head, flow, power and entrainment are presented for review. In addition, the inherent benefits of axial flow, high efficiency impellers like the A3 1 0 and A6000 in auxiliary mix chambers will be discussed. The high flow, low shear characteristics of these patented impellers provide the necessary dispersion stability for mass transfer while speeding phase disengagement in the settlers due to low power and low shear input.

Jan 1, 1995

By Erik Westman, Enji Sun, Xu Ma, Ben Fahrman

"Shale gas has become an increasingly important clean energy, which has been explored worldwide in recent decades. Supercritical CO2 acts as fracturing fluid for shale gas production. The safety monitoring is essential to prevent any kinds of leakage from the reservoir as the supercritical CO2 physically stored hundred kilometres underground. Seismic tomography is an imaging technique that uses induced seismic waves to create three dimensional images of the subsurface. It is an effective monitoring method to evaluate the caprock integrity in the carbon dioxide sequestration storage (CCS). In this experimental research, a simulated uniaxial compressive load is applied on a granite sample to analyze the stress redistribution for long-term in-situ caprock integrity during CO2 injection. The induced seismic waves are recorded and seismic events are traced based on the Geiger algorithm. The frequency of seismic events correlates with the caprock failure evolution. The acquired seismic data is divided into four regimes based on the frequency of seismic events and the failure process to examine the failure evolution. Furthermore, the travel time and distance is plotted to analyze the variation of velocity. Finally, the double difference tomography (TomoDD) algorithm using arrival time is adopted to recalculate the locations of seismic events and velocity structure in each regime. The results indicate that the passive seismic system can map the caprock stress distribution and allow for imaging of the caprock integrity. TomoDD exhibits sound improvements to relocate seismic events both in relative and absolute locations as well as to characterize the local velocity structure. The study further reveals that seismic monitoring along with TomoDD could evaluate the caprock failure accurately in the CCS.Introduction To reduce the CO2 concentration in the atmosphere, CO2 capture and sequestration (CCS) has been suggested to as a means of continuing to use fossil fuel resources while offsetting their negative environmental impacts (Eiken et al 2011). A positive CCS project should retain 99% of the injected supercritical CO2 (pressure is greater than 7.38 MPa and temperature is above 31.04? and a density is of about 700 kg/m3) over at least 100 years (Davidson et al 2005). As large quantities of supercritical carbon dioxide are being injected into full scale storage projects, caprock stress is redistributing and the leakage could occur anywhere over a wide storage area (Hou et al 2012). Acoustic waves are emitted from the caprock, as it fractures due to the CO2 injection, which can be detected by passive seismic sensors (Shitashima et al 2013). Tomography techniques are advantageous for imaging the integrity of the caprock in the CCS projects. Velocity tomography uses waves to model entities based on the arrival time of waves (Westman et al 2001, Westman 2004, Luxbacher et al 2008). Seismic tomography is a data inference technique that exploits information contained in seismic records to constrain 2D or 3D models of the Earth's interior. It generally requires the solution of a large inverse problem to obtain a heterogeneous seismic model that is consistent with field observations (Kudryavtsev et al 2012)."

Jan 1, 2017

By Oren Tranbarger

The ability to sense or detect the coal/non-coal interface during mining can provide numerous advantages for improving net coal recovery and quality, for increasing productivity, and for mining coal which would otherwise be ignored. This is particularly true when considering multiple, pitching, and thin coal seams. This paper discusses the application of coal interface sensing and examines several potential coal interface sensing techniques with respect to state-of- the-art, design, performance, and cost criteria. Four interface boundaries are considered: the roof, the floor, the rib, and the forward region.

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 T. A. Oberhellmann

The Wobbler Feeder is a comparatively new materials handling and separating device. It is essentially a materials separating or scalping machine, particularly adaptable to separation of oversize particles from mine-run ore and to the separation of oversize materials from overburden, where the removal of oversize materials is necessary, as is the case in conveyor stripping. In construction, the Wobbler Feeder is simply a set of transverse bars of manganese steel, or other wear-resisting alloy steel, each having an elliptical cross-section with the major axes of adjacent bars displaced by ninety degrees so when rotating in unison, the pre-determined space between the bars remains constant.

Jan 1, 1959

By M. E. Rossi

Reconciliation of long-term, block-model reserves and mine production frequently shows significant discrepancies. This reconciliation can be done comparing the long-term, block-model reserves against a grade control model (a model based on blast holes or production data). In this article, this is called an Fl factor for tonnages, grade and ounces. This reconciliation should also be done by comparing the grade control model to the mill feed (or loaded to heaps) tonnages, grades and ounces (F2 factor). The long-term model usually contains blocks with dimensions larger than the size of the selective mining unit (SMU) of the operation. The smoothing effect of an estimator, such as kriging, will generally result in a grade-tonnage curve that does not match the grade-tonnage curve of the SMUs to be recovered. This problem is partly to blame for the reconciliation discrepancies found. In addition, in-pit perfect selection is impossible. Therefore, the grade-tonnage predictions based on blastholes also need to be assessed - considering unplanned dilution and the errors of estimation of the grade control model. This article focuses on using conditional simulations to assess the uncertainty of the predicted grade-tonnage curve in terms of minable reserves and in the framework of long-term mine planning. Quantifying the uncertainty in no uncertain terms is done in conjunction with a popular pit optimizer. It yields several possible minable reserves and several possible minable tonnages and grades, per pit shell or scheduled phase. This is a step further from the usual sensitivity analyses that are performed for deposit evaluation or development planning. Specifically, this approach takes into account the volume-variance relationship, the selectivity of the mining operation and the planned and unplanned dilution. It also carries the local uncertainty implicit in grade estimation into the mine design and mine planning phases. The estimation of recoverable reserves revolves around estimating the block distribution of metal grades from the drill hole information available, typically composites of a fixed length. Ideally, this estimation should accurately predict the ore sent to the processing facilities of an operation. Here are some definitions used in this article: ? Recoverable reserves are those where an attempt has been made to estimate the actual distribution of grades (usually by a block model) that would be found in the ground at the time of mining, relative to the mine equipment size and other operational characteristics of the mine SMU. In particular, it incorporates the effect of internal dilution for each block how the grade distribution changes in variability when comparing a composite grade distribution to the "mining unit" grade distribution. In geostatistical jargon, this is known as "change of support effect," or volume-variance effect. ? Minable reserves are those recoverable reserves defined by a mine plan. In an open-pit scenario, they would incorporate pit optimization in addition to the sequencing and scheduling of production through mine life. ? All other block model estimates from drill hole data that do not explicitly attempt to estimate or measure the volume-variance effect are called in situ. It can be argued that these models may be inappropriate tools for predicting the grade-tonnage curve of the material to be sent for processing (minable reserves). The purpose of estimating recoverable reserves can be thought of as estimating the appropriate amount of dilution that should be incorporated into the model so that it adequately predicts (long-term and on average) the grade-tonnage curve of the ore sent to the processing facilities (mean grades and ton-

Jan 1, 1999