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By Jon Y. Kaneshiro, Stephen J. Navin, Jerry Ostberg

In 1968, the U.S. Bureau of Reclamation awarded S.S. Mullen Construction of Seattle, the Intake Tunnel (488 m by 4 m diameter tunnel) and Pumping Plant No. 1. In1998, the Southern Nevada Water Authority awarded Lake Mead Constructors (Kiewit of Omaha) a parallel New Intake Tunnel (518 m long by 4.6 m diameter tunnel) and Pumping Plant. Both projects provide water to a thirsty Las Vegas. This paper presents a detailed comparison of the two intake tunnels in terms of grout takes and pressures, geologic mapping and rock quality, grout penetration beyond the excavation line, degree of fracture grouting, and groundwater inflow boundary considerations. The paper provides a comparison of grouting techniques and material developments then and now.

Jan 1, 2001

By William Hodder, Lisa Smiley, Ehsan Alav

The Don River & Central Waterfront Coxwell Sanitary Bypass Tunnel (CBT) Project in Toronto will extend approximately 10.5 km and be lined using a precast concrete segmental tunnel liner with an internal diameter of 6,300 mm and external diameter of 6,900 mm. The CBT liner was grouted through ports drilled in the segments. The Westerly Storage Tunnel (WST) in Cleveland will extend approximately 2.93 km long and will be lined by using a similar precast concrete segmental tunnel liner with an internal diameter of 7,620 mm and external diameter of 8,230 mm. The WST liner was grouted using both grout ports built into the tail can of the TBM and through segments. Each tunnel was mined primarily in shale with varying levels of weathering and both tunnels had dry and wet sections. This paper will document the means and methods for grouting in each tunnel as well as the methods utilized to proof grout and the results found.

Jun 13, 2021

By M. Pillay

INTRODUCTION Heat stress continues to be a safety concern in underground mining. This is due to a number of reasons including increasing depths of mining, increased levels of heat generated from machines, ventilation systems being constrained from providing suitable levels of airflow and others. In order to assess if a thermal environment is safe for miners under normal or modified work arrangements, or to specify stop work conditions a heat stress index is used to guide decisions. Historically, a number of heat stress indices have been developed and applied in underground mines. These can be classed under three broad headings; single measurement, empirically based and rational indices. Do any of these indices perform better than the others? Which one allows simple determinations and decisions to be made in the mine? This paper aims to provide answers to these questions by comparing and contrasting three indices; effective temperature (ET), simple air cooling power (ACP) and Thermal work limit (TWL) for a range of typical mining wet and dry bulb combinations? The paper concludes with the identification of areas for future research. BACKGROUND Heat stress causes discomfort, increases physiological strain, decreases productivity and performance, and contributes to increased incidents [1]. Effective management and control of heat stress is particularly important in mining where hot working conditions are very common [2]. Heat stress index is a common indicator that is used by mine safety advisors and managers to guide decisions about safe work practices in mining under normal or modified conditions. According to Howes and Nixon [3] A safe heat stress control strategy for an underground mine has three elements: • Application of an environmental measure which reflects physiological strain with sufficient accuracy for the range of conditions encountered underground. • Acceptance of a functional relationship between the environmental measure and human performance which is used to optimise the environmental conditions achievable with either ventilation or ventilation and refrigeration. • A management control strategy based on the environmental measure which is designed to ensure that work in environments where excessive physiological strain may occur is prevented and corrective action is initiated. The link between these elements is an environmental measure that reflects the physiological strain a worker is subjected to. Since the turn of the twentieth century various indices of thermal stress have been developed (McPherson [4], Howes and Nixon [3], Brake [5]). In selecting a suitable environmental measure or index one problem that must be addressed is the ease with which it can be physically obtained relative to it accurately reflecting the physiological strain. No index can be universally applied and there is no single correct measure or index. However if an index has provided adequate control in a particular circumstance, then it can be regarded as correct for that situation.

Jan 1, 2019

By D. Marcotte, R. V. Dutaut

"In mineral resource evaluations, geostatistical methods known as global change of support allow prediction of the theoretical histogram and gradetonnage curves prior to interpolations or simulations of grades. Two methods commonly used by professionals to guide the choice of interpolation parameters and assess results are the discrete Gaussian model (DGM) and the indirect lognormal correction (IndLog). These models rely upon an estimate of the dispersion variance of the blocks, which is derived by numerical integration of the variogram model over a discretized block. Due to difficulties in obtaining well-formed traditional experimental variograms (especially in the presence of outliers and limited clustered data), many professionals prefer to use ‘normalized’ variograms such as correlogram (non-ergodic variogram), pairwise-relative, or variogram of the normal score transform. A series of simulations with different grade distributions and variogram models are used to assess the performances and robustness of the various variogram estimators with respect to the DGM and IndLog global change of support. Our results show that the traditional variogram, the correlogram, and normal score variogram have better performances, compared to pairwise, for both DGM and IndLog. Moreover, DGM provided better results than IndLog for the grade distributions that are not strictly lognormal. These findings provide valuable guides for geostatistics practitioners. IntroductionWhen reporting resource estimates,practitioners are required to follow generalstandards described in the NI43-101 (CIMguidelines), JORC, or SAMREC codes. One cangenerally recognize three main parts directlyrelated to the block model:•,Pre-processing: exploratory data analysis (EDA), domaining, capping, compositing, declustering, variography•,Processing: block size, neighbourhood, interpolation types (linear or nonlinear)•,Post-processing: classification, reporting.Pre-processing aims to simplify and strengthen the processing step. Ideally, domaining, capping. and compositing is aimed at defining a single homogenous population. During the processing phase, a common practice is to use a block size corresponding to the planned selective mining unit (SMU). In most precious metal deposits (gold, notably), the SMU size is frequently smaller than the recommended half data spacing (Journel and Huijbregts 1978), resulting in high estimation variance and either a high degree of smoothing if the regression slope is managed or a high degree of conditional bias if it is not. Most actual resource estimates are made using block sizes between one-quarter to one-sixth of the average data spacing, sometimes even smaller. In addition to the block size, the interpolation choices for the neighbourhood selection and interpolator (e.g. inverse distance (ID) or ordinary kriging) are often based on the Qualified Person’s (the QP) experience and some basic validation plots that often do not consider the conditional bias (e.g. trends or SWATHs plots, which are designed to compare two sets of population using a one-dimensional graph)."

Jan 1, 2018

By C. Haberfield, J. Szymakowski

Recent studies by the authors into the behaviour of jointed rock masses using direct shear testing have confirmed that joint spacing has a significant effect on rock mass strength. The rock mass failure envelopes assessed from these tests have been compared to failure envelopes obtained using the Hoek-Brown GSI criterion. Good agreement between measured and estimated GSI strength envelopes where obtained when joint spacing was considered in the assessment of GSI. Les études récentes conduites par les auteurs sur le comportement des joints rocheux employant l'essai direct de cisaillement ont confirmé que l'espacement a un effet significatif sur la force de la masse rocheuse. Les enveloppes de ruptures obtenues pour ces essais ont été comparées à l'enveloppe de rupture obtenue en utilisant le critère GSI Hoek-Brown. Une bonne corrélation a été constatée entre les enveloppes obtenues expérimentalement et les enveloppes calculées par la méthode GSI quand l'espacement des joints a été pris en considération dans l'évaluation de GSI. Kuerzliche Untersuchungen der Autoren betreffs Reaktionen von gebrochenen Felsmassen zu direkter Scherkraft, bestaetigen, dass die Abstaende von Rissen einen deutlichen Einfluss auf den Widerstand der Felsmassen hat. Die Bruchbereiche der Felsmassen die von diesen Untersuchungen entwickelt wurden, wurden verglichen mit den Bruchbereichen, die man mit der Anwendund der Hoek-Brown GSI Formulierung erhaelt. Eine gute Uebereinstimmung zwischen gemessenen und errechneten GSI Festigkeitsbereichen wurde erzielt, wenn der Abstand der Risse in der Berechnung der GSI in Betracht gezogen wurde.

Jan 1, 2003

By W. A. Hustrulid

Introduction A problem that has faced mining engineers and mining geologists for many years is how to best explore a particular prospect so that the basic decision as to whether the material present is ore or only high-grade rock can be made with some confidence. It is obvious that one could not begin a large scale mining program on the basis of tests made on a few hand samples. On the other hand, one could not be 100 percent certain as to the average grade or tonnage before beginning development, as this would require the orebody to be mined out in the sampling process. The basic objectives of the exploration program can be stated as (1) to reduce mining risk; (2) to lessen the cost of development; and (3) to increase the profits of mining. The exploration program should be terminated when these objectives are no longer being met. The construction of tunnels is a specific mining operation. In the exploration program one should gather the data needed to determine (1) the type of excavation technique(s) to be used; and (2) the rate of advance that might be expected. As in the evaluation of an orebody, with the data collected it should be possible to estimate mining costs and project life. Unfortunately, at the present time, it is not possible to do this at least when dealing with medium to hard rock types. The reason for this, it is felt, is basically the inability to identify what specific rock mass properties are of use in tunnel design and to specify how these may be obtained in an exploration program. In short, the exploration program should be designed to give information regarding both the volume (length) and 'grade' (in

Jan 1, 1997

By B. J. Arnold, F. F. Aplan

A model incorporating rate and residence time to describe the froth flotation process was used to provide data for the evaluation of various kinetic flotation models and to provide scale-up information from laboratory to industrial scale coal froth flotation. Rate constants were determined in laboratory and plant froth flotation tests and the residence time distribution was measured for a bank of four 21 ft 3 (0.6 m3 ) WEMCO cells treating Upper Freeport seam coal and a bank of four 300 ft 3 (8.5 m3 ) cells treating Lower Kittanning seam coal. It was found that a lag time for continuous flotation should be incorporated into scale up models and that a kinetic froth factor ratio of between about 2.0 and 2.8 should also be used to scale up data from laboratory to plant scale.

Jan 1, 1989

By R. Bishop, J. Baggett, J. Monsalve, N. Ripepi, A. Soni

Technology plays an ever-increasing role in improving the safety and efficiency of mining operations. Laser scanning and photogrammetry are two useful methods for capturing 3D digital representations of real world objects. While both technologies have been applied to the mining industry in numerous ways, the practical applications in an operating underground limestone mine has been tested for this paper, including for visualization and site characterization. Each technology is capable of creating highly detailed geospatial point clouds, but are all point clouds created equal? This paper presents a comparison of a laser scan and photogrammetry of a limestone pillar and addresses the strengths and limitations of each method for creating digital models of operating underground limestone mines. INTRODUCTION The United States crushed stone business is a ~$14 billion per year industry representing 1,430 companies which operate 3,782 mining operations across the 50 states. In 2017 alone, 1.33 billion tons of crushed stone was produced with 76% used predominately for road construction and repairs, but also for the production of cement, lime, chemical and agricultural uses. Over 70% of domestic crushed stone producers source their material from limestone and dolomite deposits, with currently only 82 mines (2%) operating underground (USGS, 2018). Underground mining requires a different technical skillset, and poses great challenges for production that must be both safe and economical. Therefore, there are many risks that an operation must mitigate in order to produce underground limestone. Some of the challenges in underground limestone mining involve geological hazards that are unique to the rest of the industry. The geologic features such as faults and weaknesses can be difficult to assess in a production environment (NIOSH, 1998). The consequences of failure due to improperly designed underground mining ground support systems, such as from a roof failure, are not only life threatening for the miners themselves, but can easily shut down an entire mining operation.

Jan 1, 2019

To simplify programming, image analyzers commonly measure either linear intercepts or the areas of particles and grains to determine liberation based on polished section mounts. However, the concentration of minerals by flotation is based on reagents that interact with the exposed surfaces of the minerals. Thus, it is often perceived that image analyzers should measure the perimeter of the mineral of interest in the ore particles. A comparative liberation study of thirteen samples collected from a flotation plant processing a complex base metal sulfide ore showed that the liberation of the ore minerals determined by area measurements is very similar to the liberation determined by the exposed perimeters. The liberation determined by exposed perimeters is more appropriate only for those cases where the mineral texture is so complex that it is retained in the small particles generated in conventional grinding operations.

Jan 1, 2002

By H. T. Chiwaye

Risk analysis is an important step in the design of rock slopes in open pit mining. Risk is defined as the product of the probability of slope failure and the consequences of the failure, and is generally evaluated in terms of safety and economic risk. Most of the risk analysis done at present is based on the use of limit equilibrium (LE) techniques in evaluating the probability of failure (POF) of the slopes. The approach typically makes use of full Monte Carlo simulations of the limit equilibrium models, with all uncertain variables randomly varied. The number of required simulations is generally over a thousand, at times as high as 20 000, in order to produce statistically valid results of the POF. Such an approach is clearly not practical when using numerical modelling programs dueto the high computational effort required. This paper explores the impact of using numerical modelling instead of the traditional LE techniques in evaluating the probability of slope failure. The difference in the overall assessed risk, in terms of economic impact, for the mining operation is then evaluated. With numerical models, approximate methods are used in the calculation of the probability of failure instead of full Monte Carlo simulations. This paper will use a method called the response surface methodology (RSM) for estimating the POF from numerical analyses. Simple slope models were used to verify the accuracy of the RSM method by comparing the results with those obtained from full Monte Carlo simulations. It is shown that there is good agreement between the POF values computed using full Monte Carlo simulation and those obtained using the RSM method. Finally, the use of numerical modelling in the assessment of risk is shown to bring a significant difference in the result compared with that from LE methods. One of the reasons for the difference is that LE models tend to underestimate the failure volumes and hence the consequences of slope failure.

Jan 1, 2010

By R A. Binns, S Murao

The PACMANUS and Jade deposits are two active hydrothermal fields that are forming massive sulphide deposits at the Pacific rim margin. Both are associated with felsic submarine volcanic rocks and both occur in back-arc settings. However, PACMANUS is associated with almost sediment-free dacites and lies on the crest of a volcanic ridge, whereas Jade occurs in a heavily-sedimented tectonic depression associated with a basement of pumiceous rhyolite. The PACMANUS deposit is richer in Cu and Au, and Jade in Pb. Zinc, Ag, As and Sb contents are similar. Sulfide textures at PACMANUS suggest formation of Cu and Zn sulfides by simultaneous double-diffusion involving inwards-moving seawater mixing with outwards-moving hydrothermal fluid. By contrast, at Jade there is a more complex history of formation where initial growth of barite on the sediment surface, probably associated with microbial activity, formed an impervious cap which trapped higher temperature hydrothermal solutions.  

Jan 1, 1995

By Balesh W. Konda, Andrew B. Cecala, George W. Klinowski

Methane ignitions continue to be a major concern to the longwall mining industry. In spite of numerous advances in technology over the past decade, there has not been a measurable reduction in the number of longwall face ignitions. These ignitions will continue to be a problem as deeper and gassier seams are mined, as production rates are increased through numerous improvements in longwall mining equipment, and as the output of the workforce increases due to improved management-labor relations. It is necessary to know methane liberations, emissions, and flow patterns in order to effectively deal with it and to design control technology to reduce the occurrence of longwall face ignitions. A cooperative research project between the Bureau of Mines and the Canada Centre for Mineral and Energy Technology (CANMET) has been established to address longwall methane problems. The initial direction of this work was to look at methane flow patterns during the entire longwall mining process, and more specifically to examine and compare methane levels and flow patterns for advance and retreat methods of longwall mining. Both advance and retreat longwall mining is practiced in the Sydney coal fields in Nova Scotia. In 1986, a study was performed on advancing longwall panels at the Lingan Colliery, and a similar study was recently completed at the Phalen Colliery on a retreating panel. Methane monitors were placed at various locations along the face, intake, return, and bleeder airways. Handheld methane monitors were also used upstream and downstream of the shearer mining machine. This setup allowed a determination of methane sources on the longwall panels, including: liberation from cutting with the shearer mining machine, emissions from the coal seam along the face, emissions from the gob area, and methane already in the intake air. The results from both studies are reported and the differences between the two longwall mining methods compared.

Jan 1, 1989

By E Ferrow, R Herbert, M E. Malmstr, M Fuchs

The study of microbial sulfide oxidation and oxidation rates of tailings requires the use of sterile control samples in laboratory experiments for comparison purposes. The need for sterile tailings sand is also apparent in the study of microbial communities and processes. This study compares six different methods with regard to sterilisation efficiency and investigates chemical and mineralogical changes in the samples as the result of sample treatment. In addition, the effect of sterilisation on mineral weathering rates is also studied. The efficiency of the different methods with regard to sterility is estimated by most probable numbers (MPN) of both sulfur and iron oxidising bacteria. The sterilisation methods that were compared are: washing with sterile water, washing with ethanol, applying antibiotics, repeated heating to 70¦C, autoclaving, and γ-irradiation. There are several possible drawbacks with the different methods, including the risk of an applied chemical acting as a substrate for micro-organisms in the system, and the risk of altering the tailings. The results show a substantial difference in sterilisation efficiency between the treatments. For example, autoclaving, repeated heating, ethanol and irradiation efficiently kill micro-organisms, rendering an undetectable amount of bacteria after one month of incubation, whilst washing with water exhibits virtually no microbiological effect, with a bacterial content comparable to the control. The use of antibiotics for sterilisation ranks intermediate: the lag-phase was twice as long in the antibiotics treatment compared with the controls. X-ray diffraction, optical and electron microscopy, M÷ssbauer spectroscopy, and mineral magnetism indicate no significant change in the bulk mineralogical composition. However, X-ray photoelectron spectroscopy analyses show that with some of the treatments, sulfides are partially oxidised, as indicated by the formation of sulfate and partially û oxidised sulfur species at the tailings surfaces. Aerated batch weathering experiments performed on sterilised tailings samples and untreated controls indicated that while the initial rates of element (S, Fe, Cu, Zn, Mg, Na, K, Ca) release may differ between the samples, over the course of ~100 days of experiment at pH 2 and room temperature, the rates declined toward a similar value for all samples. To conclude, there does not appear to be one perfect method for tailings sterilisation. Autoclaving and heating damage the mineral grain structure and oxidise the sulfide surfaces. Ethanol, antibiotics and rinsing greatly alter the pore water chemistry through the washing procedure. In addition, ethanol and denatured antibiotics might serve as a substrate and also, in the case of antibiotics, adhere to mineral surfaces thus obstructing chemical reactions. Irradiation seems to be the preferable method on an overall basis, although the formation of surface Fe3+ was suggested by M÷ssbauer spectroscopy.

Jan 1, 2003

By Y. C. Kim

Although the first published optimum pit limit design algorithm by Lerchs-Grossmann dates back to 1965, the majority of the open-pit mines in the U.S. has not as yet adopted such a rigorous optimization algorithm. Instead, the floating cone approach initially developed by Kennecott Copper in 1961 is still used today for various reasons. However, shortcomings with the floating cone algorithm, particularly for precious metal deposits with erratic mineral distributions, have provided a renewed interest in optimizing algorithms.

Jan 1, 1988

By M-E Boucher

While the haematite concentrates produced in Eastern Canada and the magnetite concentrates that Australia will soon be producing are different feed materials, the pellets they produce have similar physical and metallurgical properties. This paper will examine the similarities and differences in the evolution of pellet microstructure for pellets produced from magnetite and haematite. The molten slag phase generated at high temperatures has a lower melting point when in contact with magnetite compared with haematite due to the higher divalent iron content. The presence of some slag is beneficial for bonding but too much can inhibit oxidation of magnetite, although slag generation at lower temperatures reduces energy usage. In pellets produced from haematite, magnetite is generated artificially from the combustion of the carbon particles added to the pellet feed as a source of energy. Maintaining the optimum production level and mineralogy of such æsecondaryÆ magnetite, allowing for its subsequent reoxidation, is important for the quality of haematite pellets. In both magnetite and haematite pellets, balancing the slag formation and oxidation (re-oxidation for haematite pellets) of the magnetite (secondary magnetite for haematite pellets) is necessary.

Jan 1, 2007

By N Morales, G Nelis

Short-term plans must meet the long-term goals while accounting for shovel selectivity, operational space, and scheduling restrictions. There are several approaches that deal with these issues to deliver a feasible short-term schedule. This paper compares two approaches based on mathematical programming to address the short-term mining cut definition and scheduling in an integrated way. They work by scheduling extraction to comply with long-term production goals, but they differ in how mining cuts are modelled. The first approach schedules Selective Mining Units (SMUs) but aggregates them around predefined locations. The second approach does not work on SMUs; it decides directly what mining cuts are to be used from a pre-defined set of shapes. It assigns a destination and schedules the extraction for the selected mining cuts. We applied both approaches in a real case study to compare their results in terms of profit, mining cut design, expertise needed, profit, and optimisation performance.

Sep 1, 2021

By B. K. Loveday, N. V. Ramlall

A study was carried out to evaluate various batch flotation models for the recovery of minerals in a UG2 platinum ore. The major minerals in a UG2ore can be grouped as platinum group minerals, chromite, and siliceous gangue. This study also examined the entrainment of minerals during flotation. The models were ranked using statistical methods and an analysis of model-fit residuals. Entrainment parameters obtained from model fitting were evaluated for logic using a simple mineral-to-water ratio versus time plot. The foremost conclusion from the study was the importance of entrainment modelling. The measurement of water recovery increased the size of the data-set, and the inclusion of a simple entrainment model was statistically significant. The overall fit to the data was improved, and the entrainment model provided logical information on the recovery of gangue minerals that were not considered to be floatable.

Jan 1, 2015

By David Evans, David Barnes, Robert Raponi

"The Williams Mine processes, on average, 6,450 t/d of gold bearing ore using two parallel grinding circuits, leaching and C.1.P., carbon stripping, electrowinning and refining. At present, each grinding circuit consists of a SAG mill in closed circuit with a spiral classifier and a ball mill in closed circuit with two stages of cyclones.Since grinding has always constituted the largest portion of milling costs, the Williams Mine has been proactive in trying to reduce such costs through innovative changes. One project under consideration is the replacement of one spiral classifier by a vibrating screen.This paper discusses the testwork performed to evaluate the replacement of one spiral classifier by a vibrating screen, and the preliminary engineering performed to date.IntroductionThe Williams Mine, located in northwestern Ontario, approximately 40 kilometres east of Marathon, is part of the Hernia gold camp. The mine is operated by the Williams Operating Corporation and is owned by the Teck Corporation and Homestake Canada Incorporated. Operation of the mine started in October 1985 at a rate of 3,000 t/d. After a mill expansion in August 1988, throughput increased to 6,000 t/d. Production has increased steadily since the mill expansion and in 1995, an average of 6,450 t/d of ore was processed at a recovery of 95.1 %."

Jan 1, 1997

By W. Tom Witherspoon

Many Engineers are skeptical and cautious about skin friction factors for the portion of drilled piers where temporary casing is used. This has led Texas engineers in many cases to eliminate skin friction values for the cased section, which necessitates deepening the drilled shaft to compensate for the loss in the cased section. This paper has provided field test data that supports inclusion of skin friction in the temporarily cased hole section as a contributing factor in total axial capacity as comparable to open hole shafts of the same diameter. Reasons for the resultant friction values are provided including; remolding of the cased clay section and the increased pier circumference necessitated by the oversized casing required for adherence to the engineers shaft diameter specification.

Jan 1, 2009

By Samira Rashidi, Robert Dunne, Raj Rajamani

"High Pressure Grinding Rolls (HPGR) are known for more energy saving, better particle weakening and higher tonnage compared to semi-autogenous/ball mill circuits. These machines are scaled up based on the roll diameter. Therefore, studies in lab-scale HPGRs are always preferred because less weight of feed sample is required to run batch tests. The 50-hp laboratory-scale machine at the University of Utah is the first-ever lab-scale machine designed by late Prof Schönert, the inventor of HPGR. This 200x100 mm [DxL] HPGR is equipped with a robust data acquisition system, which enables running steady state tests using a feed sample as small as 6 kg to 12 kg. This paper presents the HPGR grinding study on two different categories of ores, one hard ore and the other soft. A test series has been performed to evaluate the energy consumption and the relation with size reduction on open multi-pass and locked-cycle HPGR circuits. The effect of feed size distribution and operating conditions has been investigated. Locked-cycle tests result in higher reduction ratio and higher rate of fines (100 mesh, 48 mesh and 28 mesh) generation when compared to open multi-pass tests. Data shows that scale-up trend lines stay consistent even when the circuit configuration and operating conditions are varied."

Jan 1, 2014