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

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

Jan 1, 2019

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

Jan 1, 2002

By H. T. Chiwaye

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

Jan 1, 2010

By Sean Gallagher, Rhys Llewellyn, Jay Mattos

This paper contains the results of an examination of accident, injury, employment, and production information reported to the U.S. Department of Labor's Mine Safety and Health Administration (MSHA) and provides information on some of the measures longwall mine operators are using to prevent accidents In all industries, technological advances play a key role in determining those who will survive as competitive participants in an increasingly global marketplace. While mining is among the most basic of industries, technological change is neither unfamiliar nor unwelcome to miners. Historically, technological advances in mining, particularly coal mining, have tracked closely with fluctuations in demand. In the early 1900's, coal accounted for more than three quarters of all the fuel consumed in the U.S. To meet this tremendous demand, underground cuffing machines were introduced, resulting in faster, more efficient coal extraction. Transporting coal to the surface, however, remained labor intensive. In the 1920's, nearly all underground coal was still being hand loaded. Demand for coal then declined as the price of oil, natural gas, and electricity became competitive. In the late 1930's and throughout the duration of World War II, coal demand and production was high. By 1940, more than half the coal mined was being mechanically conveyed to the surface. In 1948, the continuous mining method was introduced in the U.S., further accelerating the extraction process. By the 1950's and 1960's, however, coal production again declined as railroads converted to diesel fuel and gas and oil provided a larger share of the home heating market. Underground coal mines required even more mechanization to remain competitive as an energy source. The first longwall system was introduced in 1960. By 1995, longwalls were producing more than 189 million tons of clean coal annually, surpassing continuous mining production for the second straight year. The mining of coal and other materials has not come without a huge loss of life and considerable human suffering. More than 103,000 workers died from injuries in U.S. mines in the last 85 years. Between two and three thousand miners died each year from the turn of the century until 1931, many from underground fires and explosions. The U.S. has witnessed a tremendous drop in fatalities in underground coal mines. Credit for that is due to safety inspections and enforcement activities, better safety awareness and practices among management and labor, improved ventilation, the use of safer equipment, and increased mechanization which removed many miners from the hazards at the face and also reduced the size of the mining work force. There is little doubt that longwall mining has greatly increased underground coal mine productivity. The degree to which longwall mining has contributed to a safer work environment for underground miners is more difficult to ascertain.

By J. E. Owen

EXPERIMENTS on twenty-six drilling muds were carried out in an effort to establish a correlation between the apparent viscosity values for these muds as obtained with the Marsh funnel and the Stormer viscometer. This required a calibration of both instruments in which the density of the calibrating liquids was taken into account. Corresponding to the Marsh-funnel viscosity values that were obtained at 1500 c.c. in, 1 qt. out, equivalent Stormer viscosity values were obtained at a Stormer cylinder test speed of approximately 1100 r.p.m. Further, a linear relation between the logarithms of the Stormer cylinder test speed and the resulting apparent viscosity was found to exist for the muds tested for 200 to 1000 r.p.m. The significance of the slope of this linear relationship is discussed as it applies to the field use of drilling muds. It was concluded that Marsh-funnel viscosity measurements are expressible in absolute units that may be correlated with Stormer viscosity measurements, provided these are made in the manner described in the paper. INTRODUCTION The Marsh funnel is used almost universally to determine the viscosity of drilling muds in the field. The Stormer viscometer is commonly used to determine the viscosity of drilling muds in the laboratory. With the Marsh funnel, the time required for one quart of mud to flow out of the funnel originally filled with 1500 c.c. is often taken as the measurement of viscosity. With the Stormer instrument, drilling mud is tested at a cylinder speed of 600 r.p.m. and by means of calibration data the apparent viscosity is expressed in absolute units. Neither of these instruments is an absolute viscometer, and apparent viscosities obtained with either are influenced by the arbitrarily standardized conditions under which the measurements are made. It is disconcerting to the man in the field to be told that laboratory measurements of the viscosities of a group of muds differ widely, when he has already observed that they have identical viscosities when tested in the Marsh funnel. Drilling muds are not perfect fluids, therefore it is not surprising to find a lack of correlation between the apparent viscosity of a mud as determined with the Marsh funnel and that determined with the Stormer viscometer. The work described here was undertaken to develop a correlation between Marsh-funnel and Stormer viscosity measurements. To a considerable degree this has been accomplished. This study has resulted not only in a better understanding of the two viscosity instruments and the data obtainable from them, but has given considerable information regarding the viscosity characteristics of drilling muds. EXPERIMENTAL PROCEDURE Calibration of Harsh Funnel and Stormer Viscometer The first step in the experimental procedure was the calibration of the Marsh funnel and the Stormer viscometer with

Jan 1, 1941

By Binns RA, Murao S

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

Jan 1, 1995

By R A. Binns, S Murao

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

Jan 1, 1995

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

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

Jan 1, 1989

By David A. Summers, John Corwine, Li-King Chen

The results from fracturing plexiglas beams are combined in eight existing equations to determine the surface energy of the material. The reliability of each equation is tested, and the equation in use to determine fracture toughness is found to be the most exact. This equation is then used in determination of the surface energy of Barre Granite.

Jan 1, 1971

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

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

Jan 1, 2003

By Y. C. Kim

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

Jan 1, 1988

By Robert G. Beverly

INTRODUCTION Standards limiting the annual exposure of United States uranium miners to radon daughters were established in 1967 at 12 Working-Level-Months (WLM). The standard was reduced by a factor of three, to 4 WLM, in 1971. Currently, the standard is again being examined to determine if it should be changed. Since 1967, Union Carbide has calculated individual monthly exposures in company and contract-operated mines located on the Colorado Plateau. Although it has been possible, by extensive ventilation control measures and accurate routine sampling, to meet the current exposure standard, there are many miners whose exposures closely approach the 4 WLM standard for any given year. However, it was noted that for miners who work for any extended period of years the [average] exposure was much less than the standard. The primary purpose of this paper is to show that, in effect, any annual exposure standard to radon daughters results in a long-term exposure considerably below that standard. Further, most miners, due to their job assignments and/or employment habits, only receive a small fraction of the standard. HISTORY OF EXPOSURE STANDARDS Prior to 1967, radiation protection in uranium mines was fundamentally based on a radon daughter concentration guide. In 1960, the American Standards Association published mine and mill radiation protection standards (ASA-1960). The Colorado Department of Mines, in 1961, adoped a standard which followed the ASA Standard and provided that if concentrations exceeded 10 Working Levels (WL), the area was to be shut down until corrective action was taken; if between 3 and 10 WL, corrective action was to be initiated; between 1 and 3, additional samples were to be taken and individual exposures evaluated; and if below 1 WL, conditions were considered to be controlled. In 1967, the U.S. Department of Labor issued the first exposure standard which called originally for limiting annual exposures to 3.6 WLM but which was later changed to 12 WLM. The complicated regulatory developments leading to this standard have been described elsewhere (Beverly-1969, Rock & Walker-1970). Effective July 1, 1971, this exposure standard was lowered to 4 WLM per year, which is the current standard. Over the past year, there has been speculation about the potential risk to uranium miners working at the present standard. A recent NIOSH Study Group Report (NIOSH-1980) concluded: "There is also strong evidence that a substantial risk extends to and below 120 WLM of exposure." The 120 WLM corresponds to a miner working in uranium mines for 30 years, a rare occurrence, at an exposure rate of 4 WLM per year, an even rarer occurrence. On the other hand, the General Accounting Office, in a recent Report to the Congress (GAO-1981), was very critical of reports by NIOSH on general low-level radiation risks. The GAO recognized that”...important questions remain unanswered about the cancer risks of low-level ionizing radiation exposure;" and recommended that Congress enact legislation giving statutory authority to an interagency committee to coordinate Federal research on health effects of ionizing radiation exposure. The International Commission on Radiation Protection at its March, 1980 meeting recommended limiting the inhalation of radon daughters to 0.02 J per year, equivalent to 0.4 WL, which on an annual basis would be 4.8 WLM and noted it is common to reduce this figure by 20% for allowance in the case of uranium miners for external and/or dust exposure(Sowby-1980). This is essentially equal to the present standard of 4 WLM. As earlier uranium miner exposure studies are reevaluated, and as new studies are conducted, it is important that the relationship between regulatory standards and the resulting actual exposures be recognized. UNION CARBIDE URANIUM MINING EXPERIENCE Union Carbide started mining Colorado Plateau uranium-vanadium ores in the late 1920s for the contained vanadium values. In the early 1950s, the Atomic Energy Commission contracted Union Carbide to produce uranium at mills located in Uravan and Rifle, Colorado. The company now has over fifty years of mining experience in the area. Some mines are operated as company mines and others are operated by private mining companies under a contractual arrangement. Ventilation, sampling, and exposure calculations are carried out the same in contract mines as in company-operated mines. Data presented in this report do not differentiate between company or contract employees and include all employees who worked underground any portion of a year in Union Carbide mines from 1967 through 1980. At the peak of uranium mining activities in 1970, there were 577 miners employed at year end (285 company employees and 292 contract) and 52 mines in operation (8 company-operated and 44 contract mines). Contract mines varied from two-man operations up to 15 employees. Company mines were generally the larger operations and employed from 20 to 100 miners.

Jan 1, 1981

By M-E Boucher

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

Jan 1, 2007

By N Morales, G Nelis

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

Sep 1, 2021

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

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

Jan 1, 2015

By Behdeen Oraee-Mirzamani

Tunnel stability has an important role in the production process of an underground coal mine. There are various methods for analysing tunnel stability, such as numerical methods and analytical methods. In this paper, numerical methods (Phase2 software) have been used to determine tunnel wall displacement in a mining tunnel of the Parvade underground coal mine. The Ground Reaction Curve has also been drawn using analytical methods to determine tunnel wall displacement. The comparison of results from the numerical method and the analytical method show a noteworthy difference in the tunnel wall displacement. The displacement calculated by the numerical method shows a lower value than that of the analytical method because the numerical method is more suited to modeling the various coal and rock layers, as well as the shape of the excavations found in coal mines. The methodology used in this paper together with the results obtained from both methods can serve as useful tools for the coal mine design engineer when determining the ground support requirements.

Jan 1, 2013

By David Evans, David Barnes, Robert Raponi

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

Jan 1, 1997

By W. Tom Witherspoon

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

Jan 1, 2009

By Samira Rashidi, Robert Dunne, Raj Rajamani

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

Jan 1, 2014

By I Lipton

The existence of metalliferous mineral deposits on the deep-sea floor has been known since manganese nodules were discovered in the Arctic Kara Sea in 1868. Modern exploration to define commercially viable base and precious metal deposits began in the 1970s and gained further momentum in the last decade. High grade deposits have been delineated but investors remain cautious, many considering deep sea mining projects to be riskier than terrestrial projects. This paper compares opportunities and risks presented by terrestrial mining projects with those presented by deep sea mineral deposits (Table 1). The Solwara 1 volcanic-hosted massive Cu-Au-Ag sulphide deposit, the Atlantis II Deeps sedimentary base metal sulphide deposit, and the Mn-Ni-Cu-Co nodule deposits of the north-eastern Pacific Ocean are used as examples. These are advanced deep-sea projects with published Mineral Resource estimates, on the cusp of development. Off-shore diamond mines in Namibia are also considered. These examples provide a basis for comparing exploration and geological uncertainty, the technical risks of proposed mining systems, and the potential environmental and social impacts. CITATION: Lipton, I, 2018. A comparison of opportunity and risk in deep-sea and terrestrial mining projects, in Proceedings Complex Orebodies Conference 2018, pp 102–105 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 21, 2018