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By J. Hedderson, J. Starkey, S. Reeves, B. Hilscher

"As ore sorting becomes more common, it is important to quantify its impact on comminution circuits in order to precisely predict its effect on throughput, grind, and mill operating conditions. Since sulphides and other valuable minerals are often softer then the surrounding waste rock, the ore kept by a sorting system is typically softer than the average ore. In order to accurately forecast production, new measurements of hardness need to be performed on the new concentrated mill feed. This paper proposes a methodology to examine this effect quantitatively at an early stage of project development using the SAGDesign test.INTRODUCTION The application of ore sorting in the mining industry is one of the most promising recent technologies with the potential to cause step change in the way concentrators are designed. Most typically, ore sorters are used to sort a given ore flow for grade, either to reject waste from the mill feed, or rescue ore from the waste pile. Ore sorting technologies have been successfully applied at many mining properties in various forms and while the technology has reached or is reaching maturity in other industries, it is still in development in the mining sector. As a result, there is not yet a general consensus on the considerations that need to be made when designing a plant with ore sorting from a feasibility level above and beyond what is commonly performed as best practice. This paper will discuss those considerations and propose a methodology for integrating ore sorting into a comminution circuit at the design phase. Due to the linearly additive nature of the SAGDesign Test results, a “hardness balance” can be developed to estimate the ore hardness which would be measured in any sorting mass split using only two data points. This will allow a design engineer to make more informed decisions as to the most appropriate mass split to select in relation to the incurred operating costs and grinding circuit sizing."

Jan 1, 2019

By M. Coleman, M. E. Reed

The control of feed delivery quality to the flash furnace remains an unsolved problem. Many operations around the world still have limited control of feed quality to their flash furnace. Schenck Process and WorleyParsons continue to work towards the ‘holy grail’ of simple quality feed delivering both radial feed distribution and pulseless feed delivery. Ten years ago they joined forces to offer a system based on then Clyde Process RotoFeed system which could deliver ground level installation, online mixing of materials, pulseless feeding direct to the burner, with good radial distribution. Unfortunately, this proved to be too large a step for the Flash Smelting community and met with resistance from the process owners. Since then the community has developed a range of options some of which have been implemented over this period. Despite these improvements there are still areas of concern including safety issues associated with concentrate leaks and spills, uncontrolled variations in feed rate including massive flooding events, intermittent distribution improvement and set up of the burner continues to be trial and error and hard to repeat the success in each burner campaign. These continue to impact flash furnace performance globally on a daily basis. Following the community reticence to pursue the complete solution a number of interim steps are now available to pave the way to the ultimate feed system. The mixing of material has been installed and is delivering line blended materials on a daily basis. This paper will investigate the developments over the last 10 years and discuss the pros and cons of this incremental step process for overall furnace control. INTRODUCTION Schenck Process and WorleyParsons have continued to interact and observe changes in the flash smelting industry and to consider application of the Clyde RotoFeed now incorporated in the Schenck Process ProStreamTM range for delivery of feed directly to the concentrate or matte burner. Schenck Process pneumatic injection from the Clyde Process range, ProStreamTM, utilises two types of volumetric feeders to accurately control the feed of material into an injection pipeline. These are referred to as the RotoFeed and RotoScrew technologies and the choice of feeder type is dependent on the characteristics of the material to be handled. The RotoFeed is generally used for finer, granular and powdered materials which cannot be accurately handled with a screw such as pulverised coal and copper concentrate, whilst the RotoScrew is particularly suited to coarser lump materials, particularly abrasive of cohesive materials like zinc powder, electronic scrap and metal ashes. Both units can operate at high pressures and many systems work at 3 to 10 bar with the occasional system working at 20 or 30bar for carbon to gasifiers. For the Flash furnace the ProStreamTM units can be used to deliver each of the furnace feed materials into a single line, mixing the materials homogeneously and live to the burner.

Jan 1, 2019

By A. Gupta, F. Hrach, L. R. Mendoza, K. Flynn

ST Equipment & Technology LLC (STET) has developed a novel processing system based on tribo-electrostatic belt separation that provides the mineral processing industry a means to beneficiate fine materials with an energy-efficient and entirely dry technology. In contrast to other electrostatic separation processes that are typically limited to particles >75μm in size, the STET triboelectric belt separator is suited for separation of very fine (<1μm) to moderately coarse (500μm) particles, with very high throughput. The STET tribo-electrostatic technology has been used to process and commercially separate a wide range of industrial minerals and other dry granular powders. Here, bench-scale results are presented on the beneficiation of low-grade Fe ore fines using STET belt separation process. Bench-scale testing demonstrated the capability of the STET technology to simultaneously recover Fe and reject 2 from itabirite ore with a D50 of 60μm and ultrafine Fe ore tailings with a D50 of 20μm.The STET technology is presented as an alternative to beneficiate Fe ore fines that could not be successfully treated via traditional flowsheet circuits due to their granulometry and mineralogy. INTRODUCTION Iron ore is the fourth most common element in earth’s crust [1]. Iron is essential to steel manufacturing and therefore an essential material for global economic development [1-2]. Iron is also widely used in construction and the manufacturing of vehicles [3]. Most of iron ore resources are composed of metamorphosed banded iron formations (BIF) in which iron is commonly found in the form of oxides, hydroxides and to a lesser extent carbonates [4-5]. A particular type of iron formations with higher carbonate contents are dolomitic itabirites which are a product of the dolomitization and metamorphism of BIF deposits [6]. The largest iron ore deposits in the world can be found in Australia, China, Canada, Ukraine, India and Brazil [5]. The chemical composition of iron ores has an apparent wide range in chemical composition especially for Fe content and associated gangue minerals [1]. Major iron minerals associated with most of the iron ores are hematite, goethite, limonite and magnetite [1,5]. The main contaminants in iron ores are 2 and Al2O3 [1,5,7]. The typical silica and alumina bearing minerals present in iron ores are quartz, kaolinite, gibbsite, diaspore and corundum. Of these it is often observed that quartz is the mean silica bearing mineral and kaolinite and gibbsite are the two-main alumina bearing minerals [7]. Iron ore extraction is mainly performed through open pit mining operations, resulting in significant tailings generation [2]. The iron ore production system usually involves three stages: mining, processing and pelletizing activities. Of these, processing ensures that an adequate iron grade and chemistry is achieved prior to the pelletizing stage. Processing includes crushing, classification, milling and concentration aiming at increasing the iron content while reducing the amount of gangue minerals [1-2]. Each mineral deposit has its own unique characteristics with respect to iron and gangue bearing minerals, and therefore it requires a different concentration technique [7].

Jan 1, 2019

By F. F. Roberto

Sulfuric acid has been demonstrated to be an efficient lixiviant for in situ extraction of copper in the past. It has also been proposed that microbial production (by iron-oxidizing acidophilic bacteria) of acidic ferric solution either in situ or ex situ could provide an economic strategy to generate an effective lixiviant for secondary copper sulfides like chalcocite and covellite. Copper heap bioleaching also takes advantage of this approach. One of the major environmental challenges for in situ copper recovery projects is the containment of acidic solution within the ore deposit, and ultimate restoration of the depleted ore body to its original state. A perhaps overlooked aspect of projected site operations is the possibility that transient acidic conditions and pumping of solutions from surface reservoirs into the subsurface could stimulate the growth of iron- or sulfur-oxidizing species of acidophilic microbes in the ore body, leading to additional generation of sulfate, acid, and ferric iron. This outcome would benefit extraction of copper in mixed oxide/sulfide deposits in the near term, but could complicate restoration activities. Possible impacts of pumping acidic solutions into the subsurface on microbes resident in, or introduced into the extraction zone will be discussed in the context of our understanding of copper heap bioleaching and acid rock drainage. INTRODUCTION As high-grade deposits of metals such as copper continue to decline, conventional mining methods have evolved to mine and process larger and larger quantities of ore. What is currently defined as ore today often contains less metal than the tailings of past operations. The throughput of current-day mines also leads to tremendous amounts of waste rock and tailings that may threaten the environment and communities around the mines, as has recently been experienced by 3 major tailings disposal facility (TSF) dam failures in the past 5 years. Community acceptance of large mine operations thus becomes more difficult to achieve—loss of the social license to operate. In response, the mining industry urgently pursues innovations that can fundamentally change the way in which metals are mined. In situ recovery (ISR) of metals represents one approach to a more sustainable mining industry, by virtually eliminating the need for a pit or underground excavation, eliminating the excavation and transportation of ore to a process plant and the ensuing noise and dust that is generated, and any need for waste rock dumps or TSFs. An ISR mine could have very little visible footprint beyond a network of wells, pumps, and piping to transport pregnant solution to a recovery facility, such as a solvent extraction- electrowinning (SX-EW) plant for copper. Some solid waste streams and/or storage ponds may still be necessary to treat and store bleed streams, PLS and raffinate, or as part of groundwater restoration activities. Thus capital and operating investments would be potentially much lower than that required for a conventional mine. In theory production could also be controlled to adjust to subsurface ore grade and match market demand. A not insignificant challenge of in situ recovery is the requirement for an efficient, low-toxicity, and ideally selective lixiviant for the metal that is targeted.

Jan 1, 2019

By Euler De Souza, J. Daniel Stinnette, Julian Varaschin

"This paper outlines some of the factors that affect the airflow required for the ventilation of modern diesel-powered equipment and of electric equipment and examines how they can be applied in determining reasonable ventilation rates for Tier IV compliant diesel equipment and for electric equipment operating in mining environments. Airflow calculations for modern diesel-powered equipment and for electric equipment are more complex and must be based on a combination of factors that are unique to each mine. Methodologies for ventilation rate determinations for each case are proposed in this paper.INTRODUCTIONTraditionally, airflow requirements for production and support equipment operating in underground mines have been determined by multiplying the vehicle power by a ventilation rate that was either mandated by regulations or determined empirically from known quantities. However, in light of the drastic reductions to diesel equipment emissions mandated by the US EPA Tier IV and EURO Stage 4 regulations, there is still a great deal of uncertainty in the underground mining industry among those responsible for the ventilation of planned new mines or the expansion of current mines. The question remains, what ventilation rates should the ventilation system designer apply for modern diesel equipment?Also, given the ongoing development of grid/battery-electric equipment and active construction of fully electric mines, it has become critical to determine adequate airflow rates per brake kilowatt of electric engine power. One of the most important criteria for determining ventilation rates for electric equipment is the amount of heat they generate, as this is the most significant emission produced by electric equipment (which do not produce exhaust gases or particulates). Therefore, ventilation rates required to dissipate heat from electric equipment must be suitably determined for all electric vehicle fleets operating in underground mines."

Jan 1, 2019

By M. R. Shahan, W. R. Reed, P. Zhang, J. P. Rider

Roof bolter operators may be exposed to high respirable dust concentrations on continuous miner sections with blowing face ventilation when bolting is performed downwind of the continuous miner. One solution to reduce the high respirable dust concentrations is to use a canopy air curtain (CAC) to deliver clean air from a filtered blower fan directly to the bolter operators under the canopies. The influence of CAC installation in the airflow and dust dispersion around the location of the roof bolter operator can be evaluated by using computational fluid dynamics (CFD). This study, performed by the National Institute for Occupational Safety and Health (NIOSH), considers two scenarios: (1) a roof bolting machine in the center of the entry for installation of the fifth row of bolts from the face, and (2) a roof bolting machine positioned close to the face for the installation of the last row of bolts. In both scenarios, the bolting machine is placed in an environment which contains 6.0 mg/m3 of respirable dust and is ventilated by a blowing curtain with 3,000 cfm (1.42 m3/s) of air. This environment is used to simulate the roof bolter machine operating downstream of a continuous mining machine. Two operation positions are simulated at the same bolting location: dual drill heads in the inward position for two inside bolts; and dual drill heads in the outward position for two outside bolts. The influence of the CAC on airflows and dust dispersion is evaluated with the CAC operating at 250 cfm (0.12 m3/s). INTRODUCTION Roof bolting machines are used to drill bolt holes into the roof of an underground mine at specified patterns and to install roof bolts to secure the roof from falling. At times, the roof bolting machine is also used to install rib bolts; however, only the roof bolting operation was considered in this simulation. Though the roof bolting operation drastically improved the safety of miners from roof falls, some roof bolter operators are experiencing health problems associated with respirable dust. Based on recent statistics [1], following a low point in the late 1990s where the prevalence of coal workers’ pneumoconiosis (CWP) in miners with 25 years or more tenure was approximately 5%, CWP now exceeds 10% for miners with 25 years or more of tenure. In central Appalachia (Kentucky, Virginia, West Virginia), 20.6% of long-tenured miners have CWP and 4.5% have progressive massive fibrosis, the most severe form of CWP. CWP, also known as black lung, is an occupational respiratory disease that has no cure and can ultimately be fatal. The only way to prevent the incidence of black lung is to eliminate miners’ exposure to respirable coal mine dust. Current U.S. regulations require that the overall respirable dust concentration to be 1.5 mg/m3 or less for a full-shift sampling if respirable quartz is not present; otherwise, the standard is 10 divided by the percent quartz (not to exceed the 1.5 mg/m3 respirable dust standard) [2, 3].

Jan 1, 2019

By T. M. Klemetti, T. J. Batchler, T. Matthews

From 2013–2017, there were over 1,120 documented injuries, including fourteen fatalities, from ground falls in underground mines in the United States (MSHA 2017). The majority of these ground-fall injuries were not caused by a major roof collapse, but from falls of smaller rocks from the immediate roof. Roof screen can significantly reduce the number of these injuries and has been widely used in underground coal mines for surface control. Because of the potential of reducing ground-fall injuries, the National Institute for Occupational Safety and Health (NIOSH) is further evaluating the performance characteristics of welded-wire screen as used in underground coal mines by conducting a laboratory testing program using the Mine Roof Simulator (MRS) in NIOSH’s Bruceton Research Center Laboratory in Pittsburgh, PA. The load-displacement characteristics of an 8-ft x 12-ft area of 8-gauge welded-wire screen were evaluated using a test frame designed for testing multiple load-pull locations. In previous tests, load-pull contact was always in the center of the bolted screen section. In this study, larger pull contact areas are used to simulate roof falls. The information obtained in this and similar studies can be used to better understand the loading behavior and surface control capability of roof screen in underground mines. INTRODUCTION AND BACKGROUND The large majority of ground-fall injuries are caused by falls of smaller rocks from the immediate roof. Various controls are currently being used in mines to control this surface rock, including the use of wire roof screen. In mines where wire roof screen has been installed, injuries from rock falls have been reduced dramatically (Robertson and Hinshaw, 2001). Roof screen has the potential to prevent hundreds of injuries caused by the fall of small rocks between permanent roof supports (Compton et al., 2007). Because of this potential for reducing ground-fall injuries, the National Institute for Occupational Safety and Health (NIOSH) is evaluating the performance characteristics of welded-wire screen as used in underground coal mines by conducting a laboratory testing program in the Mine Roof Simulator (MRS) in NIOSH’s Bruceton Research Center Laboratory in Pittsburgh, PA. Previous tests to evaluate the performance characteristics of various types of screen have been conducted by the University of Alberta (Tannant, 2001). In this study, the load-displacement properties of welded-wire, chain-link, and expanded-metal mesh were measured by performing full-scale pull tests. A flat steel plate was pulled through a screen test sample that was bolted to a special test frame while the pulling force and mesh displacement were measured. These tests established the general load-displacement behavior of the screen. Peak load capacities and stiffness were determined for each screen type, showing how welded-wire mesh has a much stiffer initial loading response, whereas chain-link and expanded-metal mesh have large displacement capabilities and exhibit significant post-peak ductility.

Jan 1, 2019

By T. M. Klemetti, T. J. Batchler, T. J. Matthews

How representative are partial-panel screen tests of the in situ response of screen to roof and rib deterioration and displacements? Can a more representative procedure that tests up to two full screen panels improve the utilization and effectiveness of screen as a roof and rib support? This paper details a new full-panel screen test procedure and the initial results. The full-panel test frame was designed to test a 2.4 by 3.6-meter screened area with bolt spacing as small as 0.3-meter increments. The initial design included up to 6 screen pull points, but additional locations can be added to produce a more uniform loading. The test frame was developed to capitalize on the capabilities of the Mine Roof Simulator (MRS) at the Pittsburgh Mining Research Division (PMRD) of the National Institute for Occupational Safety and Health (NIOSH). The initial tests included a baseline test of a partial panel (1.22 by 1.22-m screen section) using the original 1.52 by 1.52-meter test frame, a 1.22 by 1.22-meter screen section test in the new full-panel test frame, and two tests of 2.4 by 3.6-meter screened area in the full-panel test frame. The 1.22 by 1.22-meter screen section tests in the full-panel test frame produced a peak load 6% higher and a yield load 30% lower than the partial-panel (original) test frame. The multiple pull points of the full-panel frame produced variable yield, peak, and intermediate loads, but on average were lower than the single pull-point tests. INTRODUCTION Roof and rib surface support has become a major component of the stability and safety of underground coal openings in the past 15 years. The use of screen, either metal or synthetic, has proven to decrease the injuries of underground coal miners. By 2005, there were 466 injuries in the United States caused by roof skin failures, with numerous mines not employing roof screen. In contrast, in 2015 there were 175 injuries caused by roof skin failures, with at least some of the reduction due to the increased use of roof screen. A continued effort to increase the utilization of screen in areas of underground coal mines where workers are traveling and working can reduce these injuries even further. To aid in this effort, the testing of various screen types under various loading and attachment conditions is required. Past testing efforts have primarily consisted of partial-panel testing in a laboratory setting. Laboratory testing of roof screen, both metal and composite, has been conducted in several ways by researchers, manufacturers, and other entities. The primary focus on the testing has been evaluating weld integrity of 1.22 by 1.22-meter screen section tests of various screen types utilized in mining (Dolinar, 2006; Dolinar, 2009; Tannant, 2001). There have also been several efforts at numerical simulation of the screen and screen/rock interactions (Murali et al., 2006; Shan et al., 2014). A few full-panel tests were performed with one pull or loading point in the center of the screen panel (Shan et al., 2014). As most previous efforts have mentioned, the only true way to evaluate the capabilities of screen is to conduct full-panel testing. Single wire strand, weld, and partial-panel (1.22 x 1.22-m) tests as well as numerical modeling only provide insight into the actual response of full-panel screen. A new test frame and procedure have been developed at PMRD to conduct various loading, anchoring, and material testing.

Jan 1, 2019

By J. M. Beck

Mineral property appraisers engaged in eminent domain actions on properties for which mining is the highest and best use must be fully competent in all phases of mineral resource estimation, mine project development, and mine operations (or retain the requisite competent assistance), as well as educated in eminent domain appraisal (USPAP Competency Rule). Failure to ensure such triggers the issue of appraisal competence and demonstrates a flagrant disregard of the myriad code(s) and guidelines to which the appraiser is subject. The cavalier assertion of some vague, undefined “latitude” in the valuation process does not exonerate the appraiser from a disregard of the USPAP Competency Rule. Lack of independence (i.e., “advocacy”) in mineral property appraisal is the usual result of a weak background in mining and/or appraisal theory. While it is readily apparent to those able to recognize it in reports and testimony, judges and juries cannot be expected to detect or recognize purposefully misleading (or even fraudulent) valuation concepts or testimony. Purposeful deception by the appraiser can lead to specious over-compensation of the property owner (thus victimizing the taxpayer); or, a failure to provide appropriate and just compensation to the property owner – yet, the purpose of an eminent domain proceeding is to leave the property owner no richer or poorer as a result of the taking. This paper highlights some of the deceptive practices encountered in litigation settings and raises issues of competency, professional ethics, and the effectiveness of the inherent enforcement capabilities within those institutions awarding professional certifications. INTRODUCTION The practice of eminent domain mineral property appraisal requires (particularly with respect to undeveloped mineral properties) that the appraiser possess the requisite level of competence in all phases of mineral resource estimation, mine project evaluation, mine development, and mine operation (through closure). To the extent the appraiser lacks such competence, it is incumbent on the appraiser not only to divulge this lack of expertise but also to seek competent assistance in the performance of the appraisal. This competence or expertise can only be obtained through a combination of appropriate and relevant technical education and direct mining industry experience. There exist a number of professional certification and/or credentialing authorities that speak to such competency. In a similar manner, eminent domain mineral property appraisal (as the term implies) also requires that the appraiser possess the requisite level of competence in the practice of eminent domain appraisal. This competence or expertise can only be obtained through appropriate and relevant education, direct knowledge of appraisal theory, particularly as it relates to the concept of eminent domain, and licensure.

Jan 1, 2019

By Carricajo T., S. Nazari

This paper discusses a unique dynamic solution for optimizing comminution area in real time. The proposed technology is a combination of model predictive controller, optimization methods and real-time digital twin for mineral processing. For grinding area, the ultimate goal of an effective control strategy is to maximize the mineral processing, reduce energy consumption and stabilized size classification. To have a continuous and effective comminution operation, a reliable control system that minimizes the need for human intervention is necessary. The implemented advanced control strategy for comminution area consists of two layers in hierarchical structure. The lower layer objective is to stabilize the process stabilization using model based predictive controller (MPC) technique. The unique MPC used in this layer is called BrainWave™ chosen for its robustness, convergence and delay handling; furthermore, the developed platform offers stable and repeatable solution and has short implementation time. The upper layer is a supervisory control, which sends the preferred process operation points to the lower layer to drive the process towards optimal performance. The goal of this layer is to optimize the process targets, as well as making decision. To increase process performance, automatic ore characteristic adaptations are implemented in this layer. A digital twin of comminution area is built in the second layer that runs amongst the process in parallel. By communicating, the process data in real time, a virtual densitometer is built that is being used in our advance process control strategy. Our control strategy has been developed, tested and implemented in various areas of a copper concentration plant. The targets for implementing advance process control are usually increased throughput, decrease energy consumption and variability reduction in particle size distribution. Overall, the results showed in different mine sites are increase throughput between 2 and 5%, specific energy consumption reduction about 15% and process variability reduction up to 60%. In this paper we discuss the results of the recent implementation of this technology on Minera Los Pelambres, Minera Collahuasi and CAP Minería. Their concentrator plants have installed control systems with the most updates optimization features, for SAG mill, ball mills and thickeners.

Jan 1, 2019

By Michael Arriagada, Ivan Arriagada

"As mines return to shafts to economically exploit deeper reserves, two methods mine owners use to minimize short term costs are: sinking shafts in phases and refurbishing old, existing shafts. While there may appear to be benefits to utilizing existing infrastructure, often this is not the case, resulting in missed deadlines and added costs. Similarly, if proper long term planning is not executed when sinking a shaft in phases, additional costs and downtime will be accrued. This paper will take an in depth look at both of these options with an eye towards optimizing costs, minimizing downtime, and facilitating construction. INTRODUCTION With the onset of mechanized mining and large capacity trucks, many mines turned away from shafts and began to use ramps to haul ore to surface. However, continual exploitation of mineral reserves located closer to surface, combined with the subsequent unearthing of additional resources at greater depths, forced mines to go deeper and deeper to extend the life and keep the operation running. In shallow mines, hauling ore by truck was an economical option. However, as mines get deeper, driving ramps takes additional time, and hauling over longer and longer distances becomes less economic, both in terms of cost and time. For this reason Mexico, in particular, has seen a jump in the number of shafts which have been put into production, as well as having a number of shafts in different stages of development and still more in the conceptual stages. Many of the shaft projects in Mexico are being designed in phases, with the hoist and equipment purchased capable of reaching the ultimate planned depth of the shaft, but with sinking operations stopping short at an intermediate depth. A loading pocket is then installed to start exploiting mineral at this depth sooner, with subsequent phases planned for a later date using either a rock pentice or bulkhead to separate the sinking operations from the active shaft above. Currently, we are directly involved in two such projects which are actively being deepened, as well as having seen various others already reach completion. If done properly with the long-term plan in mind, sinking in phases can be a very viable option, allowing for faster production while fully utilizing the equipment at various depths. However, there are many issues which can arise when sinking a shaft in phases that if not properly examined and planned for from the outset, have the potential to eliminate the gains hoped for when it was decided to sink in this manner."

Jan 1, 2019

By M. Somppi, D. Weirmeir, Z. Yamak

"Milling operations commonly employ activated carbon to capture complexed gold from lixiviants in leaching processes. As the carbon moves through pipes, pumps, vessels, kilns and other processes in the gold leaching circuit, this consumable may experience losses by various mechanisms like attrition, breakage, and rupturing. Whenever possible, capturing the fine gold-bearing carbon portion of this stream could provide value to milling operations at low capital cost and manageable financial risk. Identifying, testing, and validating the successful payback of a low-tech flocculant system is described. However, the work more closely explores a Monte Carlo risk-based methodology for understanding the potential payback of this project at various confidence levels – an approach that many improvement projects can follow where some level of uncertainty exists.INTRODUCTION Gold milling operations commonly employ activated carbon to adsorb gold from solutions. Normal handling causes the carbon particles from this material to abrade, break, rupture to a size that becomes too small for further use in the process. Operations often concentrate on capturing the fine particles because gold is associated with the carbon and this metal content can be a valuable revenue stream for toll processing. However, the small particle sizes become difficult to separate from the process and can ultimately leave the facility in the waste stream. Hence, operations often place a great deal of care in capturing as much fine carbon as possible by a variety of methods including using screens, settling tanks, and chemical methods. Retrofitting a circuit to capture more carbon fines is common but can expose the operation to financial risk. Approaching this potential improvement by leveraging a Monte Carlo methodology can be effective in understanding the risk and the associated payback of a project at various confidence levels. This method provides a strong framework for capital go-no-go decisions and the structure to validate the potential project success before implementation."

Jan 1, 2019

By B. Simonian

"The Mount Milligan mine, located in central British Columbia, began operation in September 2013 with a design nameplate capacity of 62,500 dry metric tonnes per day, using a conventional SABC grinding circuit followed by bulk sulphide flotation. Over the last few years, many changes have been implemented in order to debottleneck and optimize the comminution circuit to meet and exceed the nameplate design capacity as well as improve the metal recovery of the flotation circuit. This paper discusses changes, grouped into mechanical modifications, projects, and operational changes. The implementation of these changes as well as ongoing challenges and potential solutions will also be discussed.INTRODUCTION The Mount Milligan mine began production in the fall of 2013. It is a conventional truck-shovel open pit gold and copper mine and concentrator with a 62,500 tpd design capacity copper-gold flotation processing plant. The average head grades for the life of mine are 0.18% Cu and 0.39 g/t Au. At full capacity, the mine produces an average 400 dry tonnes per day of copper-gold concentrate containing approximately 21.5% Cu and 50-60 g/t Au. Current mine life is 21 years. This paper will discuss a selection of the operational, engineering, and administrative changes that have been made at the site since the publication of An Update on Reaching Nameplate Throughput and Continuous Improvement at Mount Milligan (Yue, 2017). PLANT FLOWSHEET CONFIGURATION Run-of-mine ore is crushed through a 60-110 gyratory crusher to a top size of seven inches. Up to 100% of the gyratory crusher product is diverted to the secondary crushing circuit by a moveable gate. The material diverted to the secondary crushing circuit is fed to a vibrating double-deck screen. The screen oversize is placed on a surge pile and reclaimed by four vibrating feeders to feed two 1250 kW standard head cone crushers in open circuit. The screen undersize is combined with the crusher discharge and returns to the conveyor feeding the coarse ore stockpile along with primary crusher product that by-passes the diverter gate."

Jan 1, 2019

By T. Donkin, B. Patterson, M. I. Guthrie, J. T. Ross

"The interest in the use of sensor-based particle ore sorting has increased significantly with the recent advances in the sorting technologies and higher capacities per unit. This technology has the potential to provide great value in processing lower grade ores from the mine whilst delivering higher grades to the mill as well as in upgrading waste rock into ore. Typically, the focus is on the amenability of the ore to ore-sorting and associated reductions in the consumption of energy, water and consumables. However, layout has a significant impact on the installed cost of a concentrator and the layout impacts from fitting an ore-sorting circuit into the concentrator plan need to be carefully considered. This paper examines layout considerations that play an important role in ore sorting projects. These include additional crushing and screening requirement, transportation of the rejected material back to the waste stockpile (mine fleet requirements) and effects on concentrator footprints.INRODUCTION Ore sorting has been practiced throughout history with an early report by Agricola in 1556. Metallic and non-metallic minerals were extracted from rocks by laboriously breaking them with hammers and picks and hand-picking selected material with higher grade off a table. The concept has not changed but the sorting technology has evolved significantly over the last 20 years. Ore sorting is usually a dry process in which a portion of the material is rejected based on mineral properties and/or grade. The methods used for sensing ore particles cover a large range of physical properties. A variety of sensors are available that differ on the discrimination technique that is used, which include: photometry, radiometry, conductivity, magnetism, microwave, X-ray transparency: X-ray transparency and luminescence and UV fluorescence (Sivamohan and Forssberg, 1991; Wotruba, 2006). There are two main ore sorting techniques: particle ore sorting and bulk ore sorting. According to Nadolski et al (2018), the potential for sorting is determined though analysis of grade heterogeneity within the orebody (greater heterogeneity represents more opportunity to add value to a project) at the scale of the intended sorting unit size."

Jan 1, 2019

By Ihsan Berk Tulu, Ted Klemetti, Deniz Tuncay

"The Analysis of Retreat Mining Pillar Stability (ARMPS) program was developed by the National Institute for Occupational Safety and Health (NIOSH) to help the U.S. coal mining industry to design safe retreat room and pillar panels. ARMPS calculates the magnitude of the in-situ and mining-induced loads by using geometrical computations and empirical rules. In particular, the program uses the “abutment angle” concept in calculating the magnitude of the abutment load on pillars adjacent to a gob.In this paper, stress measurements from U.S. and Australian mines with different overburden geologies with varying hard rock percentages were back analyzed. The results of the analyses indicated that for depths less than 650 ft, the ARMPS empirical derivation of a 21° abutment angle was supported by the case histories; however, at depths greater than 650 ft, the abutment angle was found to be significantly less than 21°. In this paper, a new equation employing the panel width to overburden depth ratio is constructed for the calculation of accurate abutment angles for deeper mining cases.The new abutment angle equation was tested using both ARMPS2010 and LaModel for the entire case history database of ARMPS2010. The new abutment angle equation to estimate the magnitude of the mining-induced loads used together with the LaModel program was found to give good classification accuracies compared to ARMPS2010 for deep cover cases.INTRODUCTIONIn the early 1990s, the Analysis of Longwall Pillar Stability (ALPS) was introduced by Mark (1990) as a chain pillar design software and was generally accepted and used by the U.S. coal mining industry. Following the success of ALPS, the National Institute for Occupational Safety and Health (NIOSH) developed the Analysis of Retreat Mining Pillar Stability (ARMPS) program for designing retreat mining pillars using a similar approach as ALPS (Mark and Chase, 1997). The Australian mining industry also recognized the success of ALPS, and Colwell, Frith, and Mark (1999) calibrated the program to Australian conditions. The ALPS and ARMPS programs draw their strengths from the large databases that are used to calibrate them (Mark, 2009). However, following the Crandall Canyon Mine collapse in 2007, NIOSH had to reconsider the pillar design criteria used in deep-cover retreat mining (Mark, 2010). The ARMPS overburden load prediction algorithm was improved to more accurately predict the loading of narrow panels with high overburden depths by implementing the pressure arch concept, and this new version is called ARMPS2010."

Jan 1, 2019

By M. Sears, K. Mohamed, J. Addis, J. Wickline, G. Rashed

To better understand the load shedding and the stress transfer on coal pillars due to room-and- pillar retreat mining in deep cover panels, researchers from the National Institute for Occupational Safety and Health (NIOSH) conducted a monitoring field study. Two sites, at overburden depths of 1,000 ft (304.8 m) and 1,500 ft (457.3 m), were selected in a room-and-pillar mine operating in Southern West Virginia. The deformation and stress changes in the roof and two adjacent pillars at each site were monitored during the retreat mining process. The monitoring results and field observations were used to calibrate large-scale FLAC3D models for each site. The calibrated models were used to assess different sizes of roof bolts installed in deep cover areas. INTRODUCTION Deep cover room-and-pillar retreat mining is an important emerging issue that will intensify in the future as shallower reserves are depleted (Mark, 2009; Chase et al., 2002). During pillar retreat mining, the likelihood of local and/or global instabilities increases due to elevated stress levels near the pillar line because of the abutment loading. In the last decade, approximately 31% of the room-and-pillar mining groundfall fatalities in the United States occurred during pillar retreat mining. Between 2013 and 2017, this proportion increased to 55%, of which 50% of these fatalities were due to falls of roof and rib that occurred during pillar recovery (MSHA, 2018). One interesting piece of information about pillar recovery fatalities was that the victim was nearly always under bolted roof. Sheared and broken #5 (5/8-in), fully grouted rebar bolt failure contributed to three of four fatal roof fall incidents that have occurred in deep cover retreat mines (NIOSH, 2010). This claim raised the following question at the study mine: Is it safe to use a 4-ft long #5 (5/8-in) rebar bolt grade 60 at a depth of cover more than 1,000 ft (304.8 m)? There was some concern about shear failure of the #5 (5/8-in) rebar bolt at a depth of cover over 1,000 ft (304.8 m). To answer the question and to reduce accidents and fatalities caused by ground falls due to room-and-pillar retreat mining in deep cover panels, the National Institute for Occupational Safety and Health (NIOSH) monitored the roof and rib deformation and the change in vertical pressure within two adjacent pillars during pillar recovery. The study was conducted at two deep cover sites in a room-and-pillar mine located in Southern West Virginia that operates in the Lower War Eagle (LWE) seam. The monitored sites were selected at 1,000 ft (304.8 m) and 1,500 ft (457.3 m) depths of cover. In this paper, site-1 represents the monitored site at 1,000 ft (304.8 m) of cover, while site-2 represents the monitored site at 1,500 ft (457.3 m) of cover. For this study, large-scale FLAC3D numerical models were calibrated based on both field observations and instrumentation results. The calibrated models were then used to compare the ground response and the induced stresses of two pillars at instrumented sites 1 and 2. The maximum lateral displacement (roof shift) obtained from the calibrated large-scale FLAC3D models was applied to a small-scale FLAC3D bolted models, the safety factor for axial stresses of the simulated roof bolts was calculated at the deep cover sites.

Jan 1, 2019

By S. Hopkins, J. Nolan, S. Gillis

"Cast in place concrete shaft linings have been used in many salt and potash shafts sunk in North America over the last 100 years. A large number of these shafts have exceeded their initial design lives and as producers look to extend the life of their operational assets, maintenance costs and shaft downtime will, in many cases, continue to rise unless repair work is undertaken. Successfully implementing concrete liner repair schemes in operating salt and potash shafts is complicated due to their age, environment and time constraints during inspection, detailed design and construction. This paper outlines a general approach to the repair of concrete-lined portions of these shafts, based on lessons learned from the authors’ involvement in a recent repair project. INTRODUCTION Prior to the recent boom in potash mine construction many of North America’s potash mine shafts were entering their sixth continuous decade of production (Table 1). Of North America’s salt mines, one has been in continuous production, for nearly nine decades.As would be expected, considering the environment in which these liners operate and the age of the material, the concrete-lined sections of the shafts outlined in Table 1 have been noted by the authors to have lost both cross-sectional thickness and strength over time. Since construction, three principal factors have contributed to this loss in section/strength; (1) physical degradation (impact, freeze/thaw or rebar expansion), (2) chemical degradation from brine/water inflows (Neville 1969; Bothwell, 1972) and (3) overstressing of the liner. These factors may act in combination, in which case they can produce more significant distress than individually. Physical degradation There are three main causes of physical degradation of concrete shaft liners. The first is impact from falling debris/tools or equipment (whether large or small), which may damage specific sections of concrete lining as they fall through the shaft. Impacts tend to damage infrastructure fixed to the shaft liner more than the liner itself, so damage from this mode of failure is usually localised, for example, where bolts fixing a shaft services bracket to the wall have been pulled away, pulling out the concrete around the bolts with them."

Jan 1, 2019

By A. Pearson, Auld. F. A., S. O. Colbeck

"To minimise mine shaft maintenance over the long-term, some potash mine shaft sinking projects have required the provision of a continuous fully welded waterproof steel liner over a substantial depth to achieve zero water ingress. This brings many challenges, including the implementation of a bottom-up construction sequence incorporating extensive waterproof welding and how to ensure the integrity of the welded liner when subject to substantial temperature fluctuations that are likely to occur with a composite steel-concrete liner placed within an artificially frozen shaft excavation. This paper highlights the key challenges and discusses multiple mitigation measures including vertical shaft movement joints and a combined structural-thermal analysis approach. INTRODUCTION Whilst the application of combined steel and concrete shaft liners using a fully welded steel liner is not new and has been used in several previous projects, most notably at Lanigan (Storck, 1968 & Roesner, 1980), recently the need to provide a continuously welded steel waterproof liner for over 800 m length of shaft for a potash mine has been considered. The need for a fully continuous welded steel liner arises from the perceived benefits of providing a fully dry shaft environment to; increase the life expectancy of the shaft; reduce routine maintenance particularly combatting salt accretion and minimising wash down practices; and improve long-term ground stability by preventing salt dissolution. Whereas previous projects have employed discontinuous sections of steel liner, when a fully continuous liner is employed longitudinal shaft strains from thermal changes result in significant deformations that must be either explicitly accommodated or resisted. Note, these thermal vertical expansion forces are more than capable of lifting the shaft collar and head frame structures should they be fixed in place. Typically, shafts are designed to resist vertical forces (collar loads and self-weight) by creating mechanical interlock with the adjacent ground. In competent rock this occurs incrementally with the construction of the shaft liner, and occasional hitch/crib structures are constructed as intermediate foundations to guarantee the transfer of vertical forces at distinct horizons. Usually this means cribs are constructed at first rock head, above the station level, and at the sump, although sometimes additional cribs are formed when entering and exiting weak strata or when a change in shaft liner construction is required."

Jan 1, 2019

By Gamal Rashed, Zorica Radakovic-Guzina, Khaled Mohamed

"Underground coal mines use mechanical bolts in addition to other types of bolts to control the rib deformation and to stabilize the yielded coal ribs. Limited research has been conducted to understand the performance of the mechanical bolts in coal ribs. Researchers from the National Institute for Occupational Safety and Health (NIOSH) conducted this work to understand the loading characteristics of mechanical bolts (stiffness and capacity) installed in coal ribs at five underground coal mines. Standard pull-out tests were performed in this study to define the loading characteristics of mechanical rib bolts. Different installation torques were applied to the tested bolts based on the strength of the coal seam. A typical tri-linear load-deformation response for mechanical bolts was obtained from these tests. It was found that the anchorage capacity depends mainly on the coal strength. Guidelines for modeling mechanical bolts have been developed using the tri-linear load-deformation response. The outcome of this research provides essential data for rib support design.INTRODUCTIONThe design of rib bolting in U.S. coal mines is based on a trial-and-error process, which is insufficient to eliminate injuries and fatalities caused by rib falls. In the last 10 years, rib failures have resulted in 17 fatalities, representing 52% of the ground-fall fatalities in underground coal mines in the United States (MSHA, 2018). Recently, NIOSH researchers surveyed 90 sites in 14 underground coal mines in the Eastern United States and found that 11 mines are bolting coal ribs, with 4 of those mines using the conventional mechanical bolts and the others using fully-grouted bolts. NIOSH researchers are developing an engineering design procedure for rib support. Pull-out test programs for mechanical and grouted rib bolts were conducted to understand the rib control mechanisms of rib bolts. Only the test results of mechanical bolts are presented in this paper.Mechanical bolt (or point-anchored bolt) is a tensioned bolt anchored down the borehole with a mechanical expansion shell or a resin-grouted medium (Peng and Tang, 1984). The mechanical anchor bolt consists of a plug, a shell anchor, a headed bar with the top end threaded, a bearing plate, a hardened washer, and occasionally a friction-reducing washer. The shell anchor can be either standard or bail type (Figure 1). The standard type shell is held in place by a nut, while the bail anchor uses a strap to hold the shell in place (Peng, 2008). The diameter of the bar is generally 16 mm (? in) for the #5 bolt or 19 mm (6/8 in) for the #6 bolt. The diameter of the hole that the bolt is inserted into must be carefully controlled, since an oversized hole can result in poor anchorage (Moss, 1971). The anchorage is obtained by applying torque to the bolt head, which in turn pulls the plug down in the shell and expands the serrated leaves against the sides of the hole-wall. In the resin-grouted anchor, the resin replaces the mechanical expansion shell at the end of the hole. The anchorage is achieved by bonding between the resin, bolt, and the hole-wall."

Jan 1, 2019

By Kalyani Devabhaktuni, Mike Minton, Mark T. Howe, Yan Zhang

During the procurement phase of a Design-Build (DB) project, design-build contractors must often commit to a fixed price based on limited or incomplete geotechnical information. This lack of information can result in an inaccurate assessment and allocation of project risk to either the contracting authority or the DB contractor. This paper presents a case study of an elevated water tank supported by mixed ground improvement techniques, including Vibro-Compaction (VC), Aggregate Piers (AP) and Compaction Grouting (CG) in East St. Louis, Illinois. An adaptive design and contracting approach provided the best value solution for a complicated set of geotechnical problems, including static bearing capacity and settlement concerns as well as liquefaction induced settlement. Close cooperation between the owner, the tank contractor, the geotechnical engineer and the ground improvement contractor was essential to produce an effective solution. Ground Improvement techniques resolved the geotechnical challenges at the site through adaptive design methods, utilizing progressive rounds of verification and real time data acquisition to achieve performance objectives. This paper details the evaluation of geotechnical criteria, the selection of ground improvement techniques, the adaptive design and implementation of ground improvements, as well as contracting methodologies used to ensure that the best value was provided. PROJECT INFORMATION In November, 2016, CB&I Storage Tank Solutions (McDermott) submitted a design/build proposal for an elevated water tank for Illinois American Water Company (IAW) in East St. Louis, Illinois. The purpose of this tank was to have backup water capacity for power generation. In the RFP, IAW requested two alternate styles of tank: Option 1, a 1250 Million Gallon (MG) Composite Elevated Tank or Option 2, a 1250 MG Waterspheroid tank. Option 1 consisted of a steel tank supported on a concrete tower. Option 2 consisted of a steel tank supported on a steel shaft. The height (determined by IAW) was 140 feet to the top capacity line measured from top of floor slab.

Jan 1, 2019