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By J. C. Salas, A. Weishaupt, J. A. Guzman

The origins of the electrorefining process date back to 1869, when the first ER cathodes were produced at Pembrey, South Wales. By the time that Pembrey was closing, Chuquicamata tank house was under construction, initiating in 1915 a long history of electrolytic cathodes production. Initially electrowon cathodes and later electrorefined cathodes were produced. During the history of Chuquicamata tank house, there have been several challenges associated with the quality demanded for electrolytic copper. Among them, the transformation from electrowinning to electrorefining in 1952, the onset of the direct commercialization of ER cathodes and the phase out of wire bars in the 80s, several technological changes including the implementation of permanent cathodes, and certainly the constant increase in complexity of the anodes refined. This paper explores the manner in which Chuquicamata has addressed these challenges and examines its links to process characteristics, which have been necessary to maintain a product quality according to market requirements and a brand, “CCC”, that is widely recognized worldwide. INTRODUCTION Located in the north of Chile, in the Atacama Desert, Chuquicamata is an open pit copper mine that for many years has played a relevant role in the global and local mining industry. Known from prehispanic times, the Chuquicamata deposit was initially developed between 1912 and 1915 by the Chile Exploration Company, the operating company of the Chile Copper Company controlled by the Guggenheim Brothers in the USA (Yeatman, 1916). Officially inaugurated on May 18th, 1915, by the President of Chile who pressed a button in Santiago sending a telegraph signal to the power generation plant specially built in Tocopilla, 1500 km north. It signaled the beginning of generating energy, sending it to Chuquicamata, another 140 km east (“El mineral de Chuquicamata”, 1915), marking the beginning of its industrial operation.

Jan 1, 2019

By W. W. Culver

The paper explores the substance, circumstances and significance of the Whelpley and Storer Copper Process. The 1866 Whelpley and Storer patent specified a process combining mechanical and chemical engineering. The process called for a down-drafting shaft furnace with a horizontal flue connected at the bottom of the shaft. Powdered sulphide ore was roasted in the shaft and then sent forward fully oxidized to a bath of sulphuric acid and/or calcium chloride. The “water furnace” was developed in Boston, at a time when the city was home to most Michigan copper mining companies. The process was field-tested at Quebec’s Harvey Hill copper mine. In 1865 James Douglas, Jr. took over management of the Harvey Hill mine with the challenge of saving the family investment. He needed a way to make money from low-grade sulphides. Through his relationship with chemist T. Sterry Hunt, Douglas learned of a possible way ahead using the process developed by Whelpley and Storer. The paper confirms the centrality of publications, meetings, and networking in the advancement of metallurgy. The process is of interest historically, as much as anything, because it is one of the pillars of James Douglas’ distinguished career in copper. INTRODUCTION At the 1866 meeting of the American Association for the Advancement of Science (AAAS) in Buffalo, New York, T. Sterry Hunt (1826–1892) read a paper “On the Metallurgical System of Messrs. Whelpley and Storer.” The paper is part of a chain of events that constitute the foundation of James Douglas’ (1837–1918) success in the copper mining industry. To understand the career of Douglas over the early years of modern copper mining, we must necessarily start with Whelpley and Storer. Going back to Hunt’s paper, he concluded his remarks by asserting that the Whelpley and Storer Process was “capable of such wide and varied application, that it is probably destined to have a very important influence on the industry of the world” (Hunt, 1867a). He reached this conclusion reasoning that the process had overcome a critical metallurgical hurdle (Hunt, 1867a). Hunt, a geo-chemist in modern terms, recalled that the chemist of those years easily separated copper from “sulphuretted” ores by converting it to a soluble salt from which pure metal was precipitated by “an equivalent of iron.” The problem was to “…apply this principle in the working of copper ores on a great scale...” (Hunt, 1867a). Hunt noted that this very problem had “…engaged the attention of many metallurgists, and has given rise to various processes for obtaining copper by the moist way, many of which are nearly or quite perfect in their operation, but are lacking in practicability and cheapness” (Hunt, 1867a). Hunt thought Whelpley and Storer had overcome the low-grade sulphide hurdle with a commercially viable engineering breakthrough. Yet in the history of copper metallurgy, the process is missing. What happened?

Jan 1, 2019

By Martin G. Wafforn, George J. Greer

"In early 2014, Pan American Silver Corp. commenced a project to construct a second shaft at its La Colorada underground mine located in Zacatecas State in Mexico. This was the first new shaft development project undertaken by Pan American Silver in the history of the company. The existing mine workings permitted the use of a raise bore for the excavation of the shaft opening, utilizing directional drilling in poor ground conditions to advance the pilot hole. This was an interesting project with many challenges and successes, and numerous lessons learned. The project was successfully completed in 2016, with the shaft operating at full mine production capacity within weeks. INTRODUCTION The La Colorada Mine has a long history dating from at least the 1920’s. The mine is located in the northern portion of Zacatecas State in Mexico in the Chalchihuites area, an approximate two hour drive south from the city of Durango. Pan American Silver acquired the La Colorada mining concessions in 1998, with mine production at that time consisting of only a few hundred tonnes per day. Over the course of the following 15 years, with the addition of an oxide processing plant and some expansions to the sulphide processing plant and site infrastructure, production was eventually increased to a maximum of 1,250 tonnes per day, all of it generated from conventional cut-and-fill stopes in the underground mine. An existing shaft was used to hoist the ore to surface, and over the years Pan American Silver had deepened the shaft and had changed out some of the hoisting plant. However, there were historic workings in close proximity to the shaft that made it unsuitable for personnel transport, and shaft maintenance was becoming difficult. The shaft was increasingly considered as a risk to the long term production continuity, and also to further deepening of the mine. In addition, expanding the capacity of this old shaft was not possible. Drilling of the La Colorada mineralization at depth commenced in 2010, with very good exploration success. As a result, by the end of 2013, the mine reserves had grown to a total of 6.5 million tonnes of high grade ore (388 g/t Ag, 0.36 g/t Au, 2.4 % Zn, and 1.3% Pb). Furthermore there were more (and higher grade) sulphide reserves than oxide reserves at depth. Following the completion of an internal study in 2013 with assistance from one of the international shaft contractors, a decision was made to increase the production from the mine with the construction of a new shaft and a new sulphide processing plant, a project with a calculated payback of 2.2 years."

Jan 1, 2019

By D. Tetsuka, H. Okamoto

The amount of impurities to be processed in a copper smelting process has increased for recent years because of an increase in the electronic scrap amount. It has been found that Sb and Ni effect on the operation of our copper electrorefining process. It is found that the current efficiency of the copper electrorefining tended to be decreased as the Ni concentration in the electrolyte increased. The passivation of the process seemed to be more related to SO4 concentration in the electrolyte rather than the Ni concentration in it. The current efficiency decreased when the Sb content in the anode was high. A high Sb content in the anode caused easy passivation by forming a gel-like film in the slime layer. We have developed a method to for see the anode passivation in a short time while generating a slime layer. We confirmed the tendency of passivation through the experiment of changing Pb content in anode. Passivation was easily occur with an increase in Pb content in the anode. The amount of slime increased as the Pb content in the anode increased. This suggests that passivation was facilitated because liquid diffusion near the anode was inhibited. INTRODUCTION The amount of impurities to be processed in copper smelting has been growing in recent years because the amount of electronic scrap has increased. Figure 1 shows that the contents of Sb and Ni in copper anodes have increased since 2007.High Sb and Ni contents in the copper anode decreased the current efficiency of the process and made the quality of copper cathode poor because of the anode passivation and the formation of floating slime in the electrolyte during electrorefining. If these influences of impurities are suppressed, it would be possible to make copper electro-refining current density higher and to process much amount of electronic scrap. We examined the components of our copper electro-refining process like anode composition, slime properties and electrolyte composition, and investigated the mechanism by which Sb and Ni affect passivation and current efficiency. When examining the influence of Ni, we noticed that the SO4 important. The effect of SO42 – was also discussed

Jan 1, 2019

By J. Gustavsson, S. Holmin, Å. Afvander, E. Rosseland, E. Zimmerman

There are numerous advantages in replacing lead-based anodes with mixed metal oxide (MMO) coated anodes in copper electrowinning; for example, lower energy consumption, improved working conditions for operators and less environmental impact. Since 2012, Permascand AB in Ljungaverk, Sweden has evaluated Permascand coated titanium anodes in copper electrowinning pilot trials at Glencore Nikkelverk in Kristiansand, Norway. The investigation has proceeded in three different steps (Phase I, II and III). In Phase I, different Permascand coatings, electrode structures and coating methods were investigated. The results showed that improved current distribution, lower energy consumption and longer service life could be obtained with expanded mesh electrode structure in a box configuration compared to rod anodes in a single-mode construction. In addition, improved current distribution and less coating wear was found when electrostatic spray coating was used instead of brush (manual) coating. It was also found that the extrapolated service life from controlled lab experiments at 200 A/m2 was much shorter compared to what was experimentally obtained in copper electrowinning pilot trials at similar current densities. This was mainly attributed to different coating failure mechanisms in lab and industrial copper production. INTRODUCTION Traditionally lead-based anodes have been used in copper in electrowinning due to lower cost. But there are many advantages in replacing lead-based anodes with mixed metal oxide (MMO) coated titanium anodes in copper electrowinning, such as lower energy consumption, higher corrosion resistance, improved working conditions for operators and less environmental impact. Still, many facilities are using lead-based anodes, although the progression of using metal oxide coated anodes in copper electrowinning has slightly increased (Robinson et al., 2012).

Jan 1, 2019

By William M. (Bill) Shaver

"The author will review the early history of shaft sinking as well as focus on the improvements made in shaft design and construction over the last fifty years (1970 to 2019) to predict how shaft sinking will change over the next fifty years and describe what the shaft sinking world might look like in 2070. INTRODUCTION It is a great honour to be one of your keynote speakers for the Fourth International Conference on Shaft Design and Construction being held in Toronto in November 2019. This is one of the opportunities we have to thank the Mining Companies of the world as well as the industries and companies we serve for their support of the mining contracting and shaft sinking business of Canada and the world at large. Many of these companies are represented here today. Without your support many of the innovations we will talk about today would not be possible. We would also like to thank all the companies that work with us in our business including all the contractors and suppliers who provide the various services we require. These include equipment manufacturers, engineering companies, project management companies and suppliers large and small who cater to our every need as quickly as possible. Without all these partners we would not be as good as we are. Many of these companies are our sponsors this week. Thank you from all of us! This is also an appropriate time to pause and think about those Shaft Sinkers we have lost and those who have been injured on our projects around the world. Please bow your head and remember the Shaft Sinkers you have worked with, who did not return home or those who returned home injured. Think about how we can continue the work of making shafts safer every day as we go forward with our projects so everyone goes home to their families safe and sound. Pause for a minute and thank you. The subject matter we have chosen to discuss today is “Shaft Sinking in the Next Five Decades”. How will the work be done, how fast will it be done, what will it cost, and why we think this may occur? It is appropriate to consider this over the next few days with this auspicious group we have assembled. Today, we will look back at all the advances that have been made over the last many years and look ahead to the future to attempt to predict what shaft sinking will look like in 2070. When I look around the room I recognize that many of us will not be around then so we need not worry about being wrong. Suffice it to say that the next 50 years will be very interesting and exciting for the shaft sinking business."

Jan 1, 2019

By A. Mazhary, T. Huff

Sensor-based ore sorting technology is a growing area of interest due to lower-grade ore deposits and higher costs of processing, among other reasons. Among the sensors currently utilized in the particle sorting industry, X-Ray Fluorescence (XRF), as the only sensor with elemental analysis capabilities, has always attracted more attention. Currently, the bench-scale amenability tests for this sensor are performed using hand-held XRF analysers. However, this testing method overestimates the performance of the XRF sorter, and therefore, the results from these tests can be misleading and are often mispresented. The work presented here was conducted on a copper mine and investigates how the sensor’s sensitivity to copper deteriorates under different testing methods. This work compares and correlates the results of an Olympus hand-held XRF analyser, static, and dynamic sorter responses with the actual grades of the rocks. The results confirmed that, the sorter response was extremely prone to blind spots. Furthermore, the results demonstrated that although the static sorter’s responses were comparable in terms of rock identification to those of the hand-held XRF analyzer and assays, the accuracy of the dynamic sorter response deteriorated both as a function of grade of the rocks as well as the potential for blind spots. INTRODUCTION Sensor-based sorting technology has been gaining increasing attention even by larger mining corporations as ore grades decrease, deposits get deeper and therefore cost of mining and processing increases. There are various types of sensors that are currently being used in the particle sorting industry such as X-Ray Transmission (XRT), Color, Electromagnetic (EM), Near-Infrared (NIR) and others. However, none of the sensors are capable of identifying and quantifying particles by the elemental content, except for X-Ray Fluorescence (XRF). XRF is a non-destructive, surficial analytical technique used to measure elemental composition of materials. X-ray fluorescence happens when a secondary x-ray (fluorescence) is released after a sample is bombarded by a primary x-ray source. From a sorting perspective, the most common and expected issues that are associated with XRF sorting are lower throughputs compared to other sensors due to particle presentation limitation as well as the need to wash the material before introducing to the sorter. However, one important issue that is mostly neglected is how the amenability tests are performed and how realistic the obtained results are. Current amenability test procedure utilises a handheld XRF analyser (such as an Olympus InnovX, as used in this work) on multiple, generally 4, sides of the rock and uses the average measured grade as a criterion for separation. While this is a widely accepted technique, it fails to consider some important factors such as exposure time and the distance between the sensor and the particle. These factors can significantly affect the sorting outcome.

Jan 1, 2019

By Sayantan Chakraborty, Kaushik Dey, SaKaushik Dey tyabrata Behera

Drilling and blasting is a popular excavation technique in the drift of underground mine. Faster completion of drift reduces the gestation period and thus longer blast pulls are attempted with longer drill holes. This compels the blasters to initiate higher charge quantity with higher confinements and ultimately results into blast-induced cracks beyond the desired periphery line of excavation. Assessment of the extent of these crack networks is difficult and often neglected. Near-field ground vibration models are often used as damage predictors (Dey, 2007, Yang et.al 1993, Holmberg et. Al 1979) but not necessarily deals with the positions of the cracks. Qualitative measurement of actual damage extent often measured using some geophysical tools, namely, P-wave velocity (Tezuka, 1995), seismic imaging (Dey, 2011), bore hole pressure monitoring (Brent and Smith, 1996), electrical resistivity (Grady and Kipp, 1993), GPR (Adams et.al, 1993) etc. Ground Penetrating Radar (GPR) is a similar geo-physical tool, which uses Electro-magnetic (EM) pulse for characterizing the medium. Similar to other geo-physical tools, GPR also comprises of a transmitter and one/multiple receivers to generate the EM pulse and receive the same after reflection/refraction from the medium(s) respectively. After its development in 1970s, GPR gained increased popularity due to its application in many important fields, namely, identification of buried objects, geological characterization, identification of geological disturbances etc. Thus, it is felt that GPR technique can be used to detect the invisible cracks inside the rockmass from the rock surface (wall and roof) itself. Understanding the same, field experimentations are carried out in an underground manganese mine in Central India, which operates on the ‘cut and fill’ method of mining at a depth of around 300m from the surface. Wedge cut blasting is carried out in the horizontal drifts of the mine using suitable blast design. A 1.6 GHz antenna was used for the GPR scanning, in which GPR transmitter transmits signals predominantly of the said frequencies. The signals get reflected back from any type of interfaces between materials in the target area, with different dielectric properties and is received by the receiver part of the ‘trans-receiver’ antenna and memorized in the Data Acquisition Unit (DAU). The data acquired is processed using a software. Ground vibration monitoring of the blasts are also carried out to develop a vibration predictor. The numerical analysis of near-field vibration predictor model results into the estimated vibration levels at the point of identified cracks. Introduction

Jan 1, 2019

By D. Stone, R. Pakalnis, B. Seymour

Most mines using backfill routinely carry out quality control (QC) testing of backfill mixes with unconfined compression test cylinders. This is true for paste operations, hydraulic fill operations, and cemented rockfill operations worldwide. However, a large number of mines simply do not prepare and test the backfill samples in a consistent manner and fail to properly analyze or interpret the test results. Conventional concrete test criteria are not applicable given the differences in the mix constituents and the fact that a large number of the backfill QC test cylinders typically fail to achieve the target strength. Standard industry practices are needed to improve the analysis and interpretation of QC test results and to provide a better means of determining if the backfill is achieving its intended target strength. This paper will attempt to lay the groundwork for standard industry practices and will address several key questions raised by operators, such as how laboratory test data compares to in-stope performance, and how many QC tests should fail before a red flag should be raised. INTRODUCTION Although paste backfill is widely used throughout the mining industry, cemented rockfill (CRF) remains as the backfill of choice in places like Nevada where very poor ground conditions mandate the use of underhand mining. The safety and stability of undercut entries beneath backfill is a function of the quality of the cemented fill, which is typically assured through routine quality control testing. In the absence of a backfill quality standard, most mines rely on concrete standards to design and interpret their backfill quality control programs. In the USA, these standards are developed and published by organizations such as the American Society of Testing and Materials (ASTM) and the American Concrete Institute (ACI). However, the application of these standards to cemented backfills can be very difficult owing to the high degree of variability in backfill mixes and the resulting variability in test results. To illustrate, the concrete industry standards denote that an acceptable (passing) test result occurs when the average strength of the duplicate test cylinders exceeds the target strength and no single test result deviates from the target strength by more than 10% (ACI 214R). Remedial measures must be taken if three consecutive tests fail. With mine backfills, it is safe to say that these criteria are rarely achievable. A further, more dramatic example is the accepted coefficient of variation for concrete QC cylinders. According to ACI 214R, the coefficient of variation (CV) should range from below 3% (excellent rating) to above 6% (poor rating). However, if this analysis method is applied to several backfill QC databases from operating mines, coefficients of variation are calculated on the order of 50%. Clearly, there is a need to develop better industry practices for assessing the quality of mine backfill and to assist mine operators with the interpretation of test results and the development of action plans in the event of undesired outcomes.

Jan 1, 2019

By D. Parks, A. Miller, M. McNinch, R. Jacksha, H. Nickerson

Each year hundreds of mine workers are involved in machinery-related accidents. Many of these accidents involve inadequate or improper use of lockout/tagout (LOTO) procedures. To mitigate the occurrence of these accidents, new safety methods are needed to monitor access to hazardous areas around operating machinery, improve documentation/monitoring of maintenance that requires shutdown of the machinery, and prevent unexpected startup or movement during machine maintenance activities. The National Institute for Occupational Safety and Health (NIOSH) is currently researching the application of Internet of Things (IoT) technologies to provide intelligent machine monitoring as part of a comprehensive LOTO program. This paper introduces NIOSH’s implementation of an IoT-based intelligent machine monitoring system to improve safety during operation and maintenance at a concrete batch plant. INTRODUCTION Of the hundreds of machinery-related accidents occurring every year in the mining industry, incidents involving stationary machinery at surface mines continue to be amongst the most frequent [1]. A National Institute for Occupational Safety and Health (NIOSH) study showed that the majority of fatal accidents involving stationary machinery at surface mines occurred at sand and gravel (38%) and stone (26%) operations. Of these accidents, entanglement in conveyor components were the most common cause of fatal accidents (48%) [2]. The same study stated that one-third of these accidents involved improper lockout/tagout (LOTO) procedures as a contributing factor. The U.S. Mine Safety and Health Administration (MSHA) acknowledged this problem, stating in a recent request for information (RFI) [3]: “Since 2007, there have been 17 fatalities related to working near or around belt conveyors, of which 76 percent were related to miners becoming entangled in belt drives, belt rollers, and discharge points. Factors that contribute to entanglement hazards include inadequate or missing guards, inadequate or an insufficient number of crossovers in strategic locations, and/or inappropriate lock out/tag out procedures. Systems that can sense a miner's presence in hazardous locations; ensure that machine guards are properly secured in place; and/or ensure machines are properly locked out and tagged out during maintenance would reduce fatalities [3].” In response to this problem, NIOSH’s Spokane Mining Research Division (SMRD) is exploring the potential application of Internet of Things (IoT) technologies to provide cost-effective intelligent machine monitoring systems for improved worker safety [4, 5]. SMRD partnered with Central Pre-mix CRH Company (Central Pre-mix) to develop and install a proof-of-concept wireless IoT solution to monitor machinery and conveyors during operation and maintenance. The primary goals of the system were to provide real-time monitoring of access points and facilitate the planning and execution of LOTO procedures

Jan 1, 2019

By C. Salcedo, J. Vallance, J. Sáez, M. Robles, J. Spangenberg, S. Rosas, L. Fontboté

The Tambomachay ore deposit (13°28'36.78"S, 71°57'35.98"W, about 6 km north of the town of Cuzco, Peru) consists of Cu hosted in arkosic red beds of the Kayra Formation (Lower Eocene). Bornite, chalcopyrite, chalcocite, covelite, digenite, malachite, and chrysocolla occur disseminated in thin layers and in veinlets. The occurrence of the copper ores in a part of the red bed sequence containing green reducing layers, the presence of organic matter in interstices between the hypogene sulfides, and the sulfur composition of the copper sulfides (δ34S values between -16.9 and -12.4‰ vs VCDT) pointing to bacterial sulfate reduction, are strong arguments to propose that mineralization was caused by copper-bearing oxidizing saline basinal fluids that precipitate copper sulfides when they meet reduced sulfur in an organic matter-rich horizon. Faults parallel to the regional Tambomachay Fault could have acted as feeders for oxidizing basinal copper-bearing fluids. Fluid migration was probably driven by tectonically-induced topography gradient. INTRODUCTION A number of red bed Cu occurrences are known in the Cuzco region in the early Eocene – early Oligocene San Jerónimo Group (Gregory, 1916; Carlotto et al., 1996, 2011; Loza et al., 2004). In the past, copper ores were extracted in several small-scale mines in the surroundings of Cuzco, including Tambomachay, Tipón, and Zurite in the Cusco province and Ushpa, Guilda, Giovanna, and Langui in the Sicuani Province. The abandoned Tambomachay Mine (13°28'36.78"S, 71°57'35.98"W), about 6 km to the north of the town of Cuzco (Romero et al., 2015; Salcedo et al., 2017, Fig. 1) was chosen as the main site for the present research aimed to better study the genesis of these copper occurrences. Results deriving from fieldwork, microscopic characterization of ore and host rock, and sulfur isotope determination suggest that the key factor for the genesis of the copper ores is that copper transported by oxidizing basinal fluids is trapped by reduced sulfur of bacteriogenic origin located in a part of the red bed sequence that shows intercalatins of green reduced horizons.

Jan 1, 2019

By John C. Stankus, Xiaoting Li

"Underground structures such as square/arch steel sets, king beam structure, etc. are widely used for the roof support in underground mines. In designing and calculating the required carrying capacities of these underground structures, how to identify the potential rock loads on the structures over long term is a challenging question for ground control engineers. Based on design practices of both surface and underground structures, the authors have designed and applied appropriate load factors in underground structural design projects to accommodate different rock types, entry/intersection dimensions, roof geologic conditions, availability of roof bolts and water, etc. Classification of load (reduction) factors for large-span underground intersections are presented for reference. Success of designed underground structures in effectively controlling roof stability over years indicates the applicability of the load factor concept for underground structural design, and its same significance as of the load factor specifications in ASCE/SEI 7-10, ANSI/AWC NDS (2015), and AASHTO LRFD Bridge Design Specifications (2012). The load factor concept will facilitate ground control engineers to appropriately utilize popular structural design software, such as RISA-3D or -2D to efficiently design and analyze underground structures.LOAD FACTORS IN STRUCTURE DESIGN APPLICATIONSIn structures, loads cause stresses, deformations, and displacements. Assessment of their effects is carried out by the methods of structural analysis with the help of load factors. Loads can be increased or decreased by load factors to meet the requirement that design strength be higher than maximum loads. For example, in designing a staircase, a dead load factor may be 1.2 times the weight of the structure, and a live load factor may be 1.6 times the maximum expected live load based on the Load and Resistance Factor Design (LRFD) method. These two “factored loads” are combined (added) to determine the “required strength” of the staircase. Some typical load (adjustment) factors used for structural design are as follows:"

Jan 1, 2019

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

"""The theory of buckling of shafts and Steel tunnel-linings"" was published by E. Amstutz in 1970 and commented on by S. Jacobsen in 1974. The authors have revised and enhanced Amstutz’s work to address Jacobsen's comments on its limitations and to extend its validity over a wider range of pipe/shaft geometries, including those commonly found in deep mine access shafts. The author has looked at Amstutz’s assessment of strengthening with flexible shear anchorages and adapted the theory of non-linear shear deformation response of bolts to develop a method of assessing increased buckling capacity utilising the full plastic shear capacity of studs. INTRODUCTION “Amstutz Buckling” refers to the buckling of pipe embedded in concrete under external uniform pressure. In terms of a mining environment, this relates directly to the design of a watertight shaft. This load case will commonly occur where a main shaft is cast against a geostatically stable primary liner and the only load applied to the final shaft is that of a uniform ground water load. Ernst Amstutz wrote his first paper on buckling of pressure-shaft steel linings in 1950 and subsequently improved his theory (Amstutz, 1970), which was then widely adopted for use within the industry. Years later, an alternative approach (Jacobsen, 1974) was proposed. Within his paper Jacobsen critiques Amstutz work, identifying some mathematical inconsistencies and highlighting limitations of the theory imposed by approximations adopted by Amstutz. Despite these limitations Amstutz work has still been used as the favoured design theory for use in the mining industry. In reviewing this body of work the author realised that with the advent of readily available computing software it was possible to adopt an iterative approach using Amstutz approximate formulae to find a solution to the original precise formulae, thus eliminating the limitations highlighted by Jacobsen The shear deformation to load characteristics of bolts has long been established (Crawford & Kulak 1968) and used for the plastic design of bolted connections in shear. This relationship is directly transferable to the design of shear studs, typical properties are well known, and actual values easily established in simple lab tests. The author has taken this relationship and modified Amstutz theory (Amstutz, 1970) for shear anchored strengthened liner buckling, to give a more efficient solution, using readily available material properties."

Jan 1, 2019

By C. Sist, J. Shannon, J. Fossenier, J. F. Adams

With the cleanest and easiest to treat copper concentrates dwindling and becoming harder to find, electrorefineries must deal with the reality of higher impurity anodes. There are well-known strategies for managing nickel, bismuth, antimony, etc. using bleeds (Ni, Bi, Sb, As) or arsenic control (Bi, Sb) or hydrometallurgical add-on circuits (IX, MRT, SX for As, Bi, Sb). However, for integrated copper producers and those producing other base metals it is often worth considering a more holistic, global approach. This paper will discuss both the well-known strategies as well as some global strategies and the role that hydrometallurgy can play in them. Examples will include atmospheric and pressure leaching, selective precipitation and impurity recycling or fixation. INTRODUCTION In the electrorefining of copper, impure copper anodes are electrically corroded, and pure copper is deposited on cathodes by passing DC current through an acidic copper sulphate solution. The anode copper impurities either settle to the cell bottom as insoluble components (anode slimes) or remain in solution with minimal deposition at the cathodes when correct operating parameters are maintained. Consideration of impurities is important in all copper electrorefineries. A typical impure anode contains 98.5–99.5% copper (Schlesinger et al., 2011). Table 1 shows the range of impurities reported for all copper anodes in the 2007 survey of operating refineries (Moats et al., 2007). Also shown is an approximate “limit” of anode compositions within which the majority of refineries operate without additional impurity removal capacity (beyond conventional liberator/bleed circuits). The column is somewhat subjective and meant to be an approximation only since there are many factors that influence impurity deportments and refinery impurity tolerances. More recent world survey data from 2013 show the averages for all impurities trending upward from 2003 to 2013 (Moats et al., 2013). With a competitive market for good concentrate and more copper recycling from electronics (which contain high levels of impurities such as nickel, lead and antimony), this trend is expected to continue through the coming years. The traditional primary method of impurity control is refinery bleed through liberator cells which are employed in the vast majority of refineries (Moats et al., 2013). These are used to control copper (primary liberators), nickel, arsenic, antimony and bismuth (secondary and tertiary liberators). Other methods of impurity control include arsenic addition (upstream in the smelter deporting to anode or directly to electrolyte) to control antimony and bismuth (Kamath et al., 2003), ion exchange or molecular recognition technology (Hur et al., 2005; Navarro et al., 2012) for antimony and bismuth. Many refineries also operate nickel evaporators to remove nickel sulphate before recycling back acid. Some use, or are contemplating using, acid purification units to recycle acid and arsenic without nickel and making nickel carbonate (Kamath et al., 2003; Sheedy et al., 2006).

Jan 1, 2019

By G. A. Gamble

"The inlet air velocity profile of any surface mine ventilation fan is an important factor when designing the connecting ductwork from the mine shaft collar to the fan. Deficient fan performance can be the result of ductwork that adds unnecessary pressure loss and/or adversely influences the fan inlet air conditions (velocity profile). The need for optimized inlet air conditions begins with discussing the process of performance testing during the development of the fan performance curve. Additional discussions include design of the shaft collar connection and the connecting ductwork. A preferred design of ductwork for parallel fan installations will be illustrated.INTRODUCTIONThe inlet air velocity profile of any surface installed mine ventilation fan is one of the most important factors to consider when designing the connecting ductwork from the mine shaft collar to the fan. Oftentimes deficient fan performance can be traced to poorly designed ductwork that adds unnecessary pressure loss and/or adversely influences the fan inlet air conditions (velocity profile). Such designs may lead to added power requirement, excessive vibration and noise, and reduced fan performance.Understanding the need for optimized inlet air conditions begins with discussing the process utilized when factory testing takes place during the development of the fan performance curve.Additional discussions include design considerations for the shaft collar connection, the connecting ductwork, and fan inlet. Illustrations of cost savings and pressure losses will be included for audience consideration. The design of ductwork for parallel fan installations on the surface is another issue that requires consideration. A preferred method will be illustrated.The selection of accessory components, such as self-closing doors, can also add to the potential for reduced fan performance. Several components will be discussed.An elementary presentation of the Air Movement and Control Association (AMCA) Standard 201 “Fans and Systems” will be included to provide background for those audience members that want additional technical explanations of the need for a uniform inlet air velocity profile at the fan inlet."

Jan 1, 2019

By Ihsan Berk Tulu, Ted Klemetti, Deniz Tuncay, Cicek Sena, Zhao Haochen

"Analysis of Longwall Pillar Stability (ALPS) and Analysis of Retreat Mine Pillar Stability (ARMPS) treat the pillar as a passive structure that is designed to carry overburden dead-weight. This dead-weight is calculated by simple rules based on the geometry of the mining, such as “tributary area theory,” “pressure arch theory,” and “abutment angle theory.” Although the pressure arch loading approach indirectly accounts for the generally stiffer overburden response of narrow and deep panels, it does not include the effect of specific geology of the overburden in load calculations. The relationship between successful pillar layouts and overburden geology can be incorporated into the load calculation by using the Ground Response Curve (GRC) approach.This paper introduces the GRC modeling methodology to investigate the effect of overburden geology and excavation geometry on load transfer mechanisms using seven field measurement case studies from four U.S. mines. It was also shown that the modeling methodology used to derive GRC in this study approximates mining-induced stresses and deformations within 5% of the values measured in the field.INTRODUCTIONFrith and Reed (2017) stated that current state-of-the-art pillar design methods ignore the overburden mechanics and use estimated dead-weight of overburden to compute pillar sizes. In the United States and around the world, overburden loading is typically estimated by simple geometric rules, generally with the tributary area theory (TAT). The specific overburden mechanics, structural competence of the overburden strata, geology, in situ stresses, and overburden/pillar interactions are typically ignored. However, these important mechanical responses affect the stability of the mine pillars and openings; therefore, they also affect the safety of the mine workers.Research conducted by the National Institute for Occupational Safety and Health (NIOSH) after the Crandall Canyon mine disaster (Esterhuizen, Mark, and Murphy, 2010) showed that small panel width-to-depth ratios and stiff-strong overburden result in a reduction of the observed pillar loads, much smaller than the TAT estimate. In fact, as a response to the Crandall Canyon disaster, Mark (2010) implemented a pressure arch loading approach into the latest ARMPS program to recognize the inherently greater stability of narrow panels at depth. The pressure arch approach is a first logical step to introducing non-tributary area loading into the practical ARMPS program. Although the pressure arch loading indirectly accounts for the generally stiffer overburden response of narrow and deep panels, it does not include the effect of specific geology and mechanical response of overburden in the load calculations."

Jan 1, 2019

By J. Gonzales

For this paper, blending is conceived as a linear process that takes a proportion from one stockpile or location and mixes it with another proportion from another stockpile or location, with the aim of obtaining a set of attributes in the final combination that satisfies a given quality standard. A blending class is the combination of ore attributes and grade limits that should be stockpiled in order to achieve the target quality for the process Plant. Current literature considers the number of blending classes in the Open Pit Scheduling Problem as fixed; however, this paper shows how to define specific blending classes for a given deposit so that the schedule can be solved with Sequential Linear Programming. Finally, a practical application in a Peruvian copper mine is shown, involving four ore attributes: alteration code, arsenic, oxide content and copper grade. INTRODUCTION In copper ore deposits, blending for quality has not been a typical issue as it was for coal and iron ore; however, current copper deposits in the world are becoming geologically and chemically more complex than in the past and is not uncommon to have quality standards involving up to four attributes, introducing an additional complexity to the mine scheduling process [1]. In order to solve the Open Pit Scheduling Problem with multiple ore quality constraints, two assumptions are necessary [2]: - Grade constraints are defined only by specifying an upper and lower limit; - Ore attributes in stockpiles are pre-determined, i.e., the upper and/or lower grade limits defining each stockpile are already known at the time of setting the schedule. Assuming each ore attribute is classified as high, medium or low, and there are four attributes to be controlled in the target ore quality, the number of possible blending classes would be 3^4 = 81. This number is impractical to manage in Operations, and increases the complexity at the time of running the schedule; therefore, a more efficient way of classifying ore is required. Note that in practice, one blending class can be split into more than one stockpile location due to physical capacity constraints. Therefore, the number of stockpiles could even be more than 81. BLENDING CLASSES DETERMINATION For a given set of ore grades with lower and upper limits to be achieved, one has to find all feasible ore classes such that when blended, their grades will lie within those limits.

Jan 1, 2019

By Joël P. T. Kapusta, Liam J. Watt

The last extended copper smelting and converting survey dates back to the 2003 International Copper Conference held in Santiago, Chile. The following year, the Minerals, Metals, and Materials Society of AIME (TMS) started a new series in its Journal of Metals (JOM) called JOM World Nonferrous Smelters Survey devoting its first Part to Copper. The copper smelting landscape has significantly changed in the past 15 years, both in geographical terms with the shift of copper smelting production from the Americas to China and in technological terms with the growing adoption rate of Top-Submerged Lancing (TSL) technologies and the development and implementation of Bottom-Blowing technologies. A copper smelting survey update, even if not as in-depth as the 2003 Copper Smelting Survey, is therefore in order. The authors will review the latest trends in copper smelting technologies and production worldwide in comparison to the landscape of the early 2000s. INTRODUCTION This 2019 copper smelting survey is different from previous surveys as our data does not originate exclusively from responses to a questionnaire sent to all the copper smelters. Our data comes from several sources. We are therefore very thankful to the International Copper Study Group (ICSG) and Wood Mackenzie for sharing some of their information and to ENFI, Outotec, and Glencore Technologies for helping us cross-check our numbers for the Chinese smelters, flash smelting, Ausmelt, and Isasmelt technologies. Our survey presents data reporting from a total of 96 operating copper smelters from around the globe, which is significantly higher than the 50 surveyed in 2003. Although secondary smelting numbers have steadily increased over time, this paper focuses on primary smelting operations with copper concentrate as feed material. The list in this survey is still far from complete, with missing information from several known smelting operations. The total primary copper production tonnage of the smelters we surveyed amounts to 18.3 million metric tonnes from 29 countries representing 92% and 94% of total world primary production for 2018 estimated by ICSG and Wood Mackenzie, respectively. This representation has increased from 2003 when the total was 9.71 million metric tonnes from 23 countries accounting for 73% of world production. The trends presented and analyzed within this paper have stronger significance because of the increase in production accounted for. The Chinese and Russian smelter information collected is more detailed than years past. Our complete dataset is summarized in Table 1 at the end of our paper.

Jan 1, 2019

By Xinzhi (Thomas) Du

"A longwall mining method is a productive mining method of coal extraction that involves the complete removal of large, rectangular panels of coal. Since this method causes deformations through the overlying strata to the ground surface, the surface- and ground-water regime above the longwall panels may be adversely affected and drained into the lower strata. The purpose of this study is to determine the effects of longwall mining subsidence on surface- and ground-water systems.This study is a site-specific hydrogeologic analysis conducted before, during, and after panel retreat at a site located over two selected longwall panels in the Pittsburgh coal seam in Greene County of Pennsylvania. Mine subsidence causes fractures in the overlying strata and enlarges bedding plane separations, increasing the effective porosity and storativity of the overlying strata. Generally, the water in deeper aquifers located in the lower fractured zone drains into the mine void due to the vertical fractures. Moreover, the water level does not recover after longwall mining (but may recover when the mine is flooded). The upper aquifers over the thick aquitard layers may maintain the water level or drop slightly during longwall mining and recover very quickly. The reason is that the underlying aquiclude layer has less change in porosity and permeability after longwall mining (Carver, 1994).A groundwater flow model-MODFLOW was used to predict the water table contours for the periods of pre- and post-mining conditions. The hydraulic gradients of post-subsidence became larger outside the panels and smaller inside the panels than that of the pre-subsidence. The groundwater hydraulic heads above the longwall panel dropped a maximum 50 ft in the study panels. The shallows monitoring wells that were less than 50 ft deep from the surface were deeply affected by longwall mining subsidence. For the intermediate wells that were more than 85 ft deep from the surface, where strata movement acted as a unit during longwall mining subsidence, there were little impacts of longwall mining on the intermediate wells in the study area. For the deep wells that were located below the Waynesburg sandstone, the water was drained to the underlying strata or even into the mine gob."

Jan 1, 2019

By T. Braun

In 1911, a young mining engineer named Bartlett Thane assembled a large claim block, 4 miles southeast of Juneau, Alaska. Across the Gastineau Channel on Douglas Island, the Alaska Treadwell Gold Mining Company had, by 1904, produced gold valued at more than three times the purchase price of Alaska in 1867 (Kelly, 2010). Thane knew his deposit was lower grade than other mines or prospects in the area; however, if the deposit was developed on a large-scale, the trend and size of the deposit made the project potentially economic. In 1911, Thane developed a business plan for the project and successfully attracted investor interest. In 1912, Daniel C Jackling visited the property at the request of an east coast investment group. Upon inspection of 5 years of production records for the Perseverance Mine, Jackling saw an opportunity to apply the low-grade/high volume milling technology which was widely adopted to the copper porphyry deposits of the southwestern U.S. With a design capacity of 6,000 tons per day (5,454 tonnes/day), Jackling’s innovative mill design would require high volume underground mining methods and the power of two new hydroelectric projects. By 1915, the Alaska Gastineau Mining Company was in full production. Between 1915 and 1919, the mine and mill set production records for haulage and throughput. The hydroelectric projects proved to be successful engineering marvels. Operational difficulties related to ore grade, refractory elements and labor shortages interrupted production on an increasing basis. In 1921, the mining company permanently shut down operations. INTRODUCTION The Juneau Gold Belt begins at Windham Bay near the southern end of Stephens Passage and extends approximately 100 miles north to Berners Bay (Figure 1). The first large-scale and sustained underground mine, operated by the Alaska Treadwell Gold Mining Company in the late 1880’s, established the Juneau-Douglas area as the regional hub for transportation and development. In the mid-1890’s, the combined population of Juneau and Douglas reached approximately 3,000 residents, largely driven by employment and support demands of several small to mid-size underground gold mines in the region. By 1904, the value of the cumulative gold production of the Alaska Treadwell was five times greater than the purchase price of Alaska in 1867 (Kelly, 2010). By the turn of the 20th century, economic growth in the Juneau-Douglas area relied on supplies delivered from the lower 48. Passenger and freight shipping lines required 3 to 4 days to make the one-way trip from Seattle to Juneau. Established shipping routes required experienced crews due to navigation hazards and adverse weather conditions. Even with precautions at the time, steam ships and freighter vessels experienced breakdowns or encounters with natural hazards that resulted in damage or in some cases, sinking. Interestingly, with no roads or railroads connecting Juneau to other communities in southeast Alaska, Juneau still relies on seaborn shipping in the 21st century.

Jan 1, 2019