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By K. A. Grace

K.A. Grace's article advocates usage of the terms describing resources as defined by the US Bureau of Mines (USBM) and the US Geological Survey (USGS). Originally recommended in 1943, the terms were revised in 1976 and again in 1980. Mr. Grace further recommends that the term reserves should be severely restricted and that it should encompass mining and processing factors. The list of references quoted in the article does not, however, mention the study concerning reserves that was made over a period of almost four years by a joint Committee on Definitions. This committee was sponsored by three societies - the American Institute of Mining, Metallurgical, and Petroleum Engineers (AIME), the Society of Economic Geologists (SEG), and the American Institute of Professional Geologists (AIPG). Most of the detailed work was done by a working party appointed by the Society of Mining Engineers (SME) of AIME. After studying previous definitions in the US as well as in the United Kingdom, Canada, Australia, France, and other European countries, the following definitions were proposed and accepted in 1974 by the SME board of directors. (Ore reserve definitions in Rhodesia, South Africa, or New Zealand were not mentioned as having been reviewed.) Table 1 lists definitions from MINING ENGINEERING, July 1975. Notice in Table 1 it was mentioned, under Resources, that the definitions of reserves were recommended for the purposes of commercial evaluation. It was further stated that the terminology adopted by the USBM and the USGS, as released to the public, covering resources, was recognized as such. Mr. Grace also recommends that reserve calculations should allow for future losses in ore tonnage, ore grade, milling, and smelting. Unfortunately, it might take years to develop a prospect after it has been discovered. Mining methods have to be tested before being finalized in order to know what tonnage of ore is going to be lost, or how much dilution is likely to occur. The same problems are present in concentrating or in smelting. Operating mines usually know how mined grade compares with ore reserve grade. They can usually predict fairly closely what their final metal recoveries are probably going to be. It is, however, expecting too much to have to delay quoting reserves from say a recently drilled prospect. Any responsible mining geologist or mining engineer (or company for that matter) should state several things about an ore reserve calculation: that it is or is not a preliminary estimate only; that estimated grades have not been diluted to allow for a future, as yet unknown, mining method; that metal recoveries are or are not only tentative, based perhaps on drill hole samples only or on no metallurgical testing at all. Reserves of a partially explored prospect are always quoted when it becomes necessary to raise capital. If they are described as preliminary only (really just tonnage and grade estimates and not true "reserves") in addition to being accompanied by some of the items just mentioned, nobody is going to be misled. If, on the other hand, a report by a member of the SME or SEG is intentionally misleading, the individual must be reported and eventually disbarred from his or her society. To summarize, let us stay with the definitions of ore, ore reserves, and related terms as recommended and accepted by SME. I might add that on occasion I have used a fourth category in my ore reserve calculations, namely Geologically Possible Ore. This category is more or less the same as Inferred ore defined by the USBM and USGS. I have never given a grade to this category, which may be based only on geological mapping, geochemical sampling, or geophysical traversing.

Jan 8, 1985

By M. J. Allan

"The paper describes the author's experiences in the startup of autogenous and semiautogenous grinding circuits gained over the last twenty years on both large and small grinding circuits. Typical phases in commissioning from construction to startup with ore are covered. The paper also describes experiences with material handling systems, slurry pumping, vibrating screens, trommel screens and pebbles crushing systems. INTRODUCTIONThe commissioning of any new mineral processing plant is an important and exciting milestone in any project. It is at this point in time that the efforts of developers, design engineers, equipment suppliers and operators culminate in a finished product. The successful completion of any project requires that the plant achieve design throughput as quickly as possible. The success of any startup can generally be anticipated by the amount of careful planning which takes place before equipment is operated. The first part of this paper describes the steps in going from construction through to introduction of ore. The organization and planning of startup activities is also covered.The second part of this paper describes some typical problem areas and past experiences in the startup of semi-autogenous grinding circuits. Semi-autogenous grinding circuits present some special challenges in the startup phase. The description of some of these problems and experiences will hopefully provide other operators with suggestions on how to avoid (or at least prepare for) these problems."

Jan 1, 1992

By M. H. Weinecke, B. P. Faulkner

This paper will discuss the process for making lightweight aggregate, the waste materials that make good lightweight aggregate, the chemistry of making lightweight aggregate and finally a discussion of the commercial prospects for making lightweight aggregate from wastes. Svedala Industries has been making lightweight aggregate (LWA) from waste materials for the last ten years. Many clay-like materials, such as coal-fired boiler ash and washings from a granite quarry, can be used alone or with additional waste materials to produce a commercially saleable LWA.

Jan 1, 2001

Jan 1, 1967

By Mike Amick

As virgin coal reserves dwindle, more and more mines are being opened in areas with previous mining. As addressed in this conference before, there are solutions available for designing mines extracting coal from reserves with known underground workings above or below the target seam. Often times, however, mining begins with limited knowledge of the underground workings. Even with careful planning, the proposed design has limited success because the ?facts? as presented during planning turn out not to be as accurate as one would like. At National Coal?s Mine #14, mining the Walnut Mountain seam, after driving the main entries in some 34 crosscuts, it was found during routine mine examinations that the pillars had separated from the roof in a few areas of the mains. Immediate roof conditions proved to be sound but questions arose concerning the stability of the main roof and floor. Core tests were done to better determine the rock strengths and characteristics. The paper presents the ideology used to tackle the effects of mining over old works in an operating coal mine. This paper discusses the analysis of the symptoms observed and the steps taken to ensure that this pillar separation was caused by the underlying old workings and that the separation was expected and controllable.

Jan 1, 2009

By K. Takahashi, I. Iwasaki

it is well known1 that racemic mixtures of optical isomers can be separated by passing their solutions through a column packed with r- or 1-quartz powder so that one of the isomers becomes adsorbed selectively on the asymmetric quartz surfaces. Although the natural occurrence of r- and 1-quartz is more or less equally distributed, some localities show enough deviations toward a higher proportion of l-quartz2 to suggest that optical isomerism might influence quartz flotation. Beta-amines have an asymmetric carbon atom, and therefore optical activity may be expected. Beta-amines are reported to behave quite differently from alpha-amines in the flotation upgrading of certain oxidized iron ores.3 It is of interest, for both practical and theoretical reasons, to determine whether or not optical isomerism plays a role in the amine flotation of quartz. In this article the results of Hallimond tube flotation tests and of adsorption measurements using d-beta-phenylisopropylamine on both r- and 1-quartz are compared. D-beta-phenylisopropylamine was chosen for the tests because it is commercially available in optical isomers. MATERIALS AND METHODS Brazilian quartz crystals were hand-sorted according to the difference in trigonal trapezohedral faces, and each crystal was further checked by the Airy's spiral.4 The respective crystals were then crushed, and the -150 + 200-mesh fractions were screened out for the Hallimond tube tests. The samples were cleaned by brief leaching with warm, dilute hydrochloric acid, and then by repeated rinsing with distilled water. In preparation for the adsorption measurements, they were then dry-ground in a pebble mill for several hours, washed a few times with distilled water, and finally dried in a vacuum desiccator. By means of the Krypton gas adsorption method, the specific surfaces of r- and 1-quartz samples were determined to be 27,700 and 19,800 sq cm per gm respectively, the difference being due mainly to the decantation procedure rather than to the inherent difference in their grindability. D-beta-phenylisopropylamine was the only collector used in the study. It was solubilized by neutralizing with 110% of the stoichiometric requirement of 0.1 N hydrochloric acid. The specific rotation of the d-beta-phenylisopropylamine was determined with a polari-meter using sodium D line to be, Although the d-beta-phenylis~prop~lamine showed absorption maxima in the wave length region of 250 to 270 millimicrons, due presumably to the phenyl group, these peaks were too sensitive to be used for colorimetric determination of the residual concentration in the adsorption measurements. The amine-picrate method, used in the colorimetric analysis of dodecylammonium chloride,' was found to work satisfactorily for the d-beta-phenylisopropylamine. The adsorption densities were thus determined after equilibrating 40 gm of the ground quartz sample with 300 ml of solution by tumbling in separate glass-stoppered Erlenmeyer flasks for one hour. EXPERIMENTAL RESULTS Fig. 1 presents the adsorption densities of d-beta-phenylisopropylamine on the r- and 1-quarts in the pH range of 6 to 7.5. In agreement with data reported in the literature, the adsorption density of d-beta-phenylisopropylamine was slightly higher on the r-quartz than on the 1-quartz, although the difference was not as pronounced as one might expect from the capacity of r-quartz and 1-quartz to separate racemic

Jan 1, 1969

By J. A. McCamley

Gypsum scrap generated from wallboard manufacturing, construction and deconstruction activities in North America, Europe and Japan is a recognized environmental problem. Landfill disposal of this gypsum scrap is problematic, with deposits often creating hydrogen sulphide emissions and metallic sulphide groundwater leachates. Of these regions, Europe is the most advanced in dealing with the issue having enacted legislation, which will come into effect in July 2005 that will strongly encourage gypsum recycling throughout the EU. An estimated 10% to 17% of all gypsum used in the wallboard industry becomes recyclable scrap, which represents approximately 1% of the total waste stream in North America. In the United States, 2.5 to 4.5 Mt of gypsum scrap is generated each year; the volume is similar in Europe (the percentage of total tonnage is less because of the higher use of brick and concrete). The Greater Vancouver region of British Columbia has banned gypsum from its landfills, and as a result, all gypsum scrap is now recycled. New West Gypsum Recycling, a Canadian firm, uses its proprietary technology to process large quantities of gypsum scrap that is reincorporated into new wallboard, at percentages ranging up to 25%. The operation of New West Gypsum Recycling Inc. of Langley, Vancouver is examined as a case study and recommendations are provided as to how other urban regions can implement similar gypsum scrap recycling programs.

Jan 1, 2004

By Rudolph G. Wuerker

THE Krupp-Renn Process,[1] has been successfully used to treat low-grade iron ores, laterites, titaniferous sands, and other minerals, and before World War I1 25 units were built by the Krupp-Grusonwerk, Magdeburg, Germany. J. J. Howard[2] has shown theoretical advantages of the pyro-metallurgical treatment over conventional approaches; and the process was described by R. Durrer[3].

Jan 1, 1951

By S. Vivek

An atomic force microscope (AFM) was used to measure the surface forces between silica surfaces immersed in solutions of hydrolysable cations and/or sodium oleate. Hydrophobic forces were observed at the pH range where the concentration of the singly-charged hydroxo-complexes (CaOH+ or MgOH+) in solution is the highest. In the absence of oleate, the adsorption of the hydroxo-complexes was indicated by the disappearance of the repulsive hydration force. These results suggest that the hydrolysable cations adsorb on silica as singly-charged hydroxo-complexes. Based on the results obtained in the present work, a model for the adsorption of hydrolysable cations has been developed.

Jan 1, 2009

Improving the Sustainability of Primary Metal Production - The Need for a Life Cycle Approach

Sep 13, 2010

By Warren E. Wilson

THE transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,1 of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer3 concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. ENERGY. REQUIREMENTS Carrying out the idea contained in Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: [Solids load in terms of mass per unit time by W Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L Slope of energy grade line by s Mass per unit volume of solids by m, Mass per unit volume of liquid by ml Mass per unit volume of mixture by m,,, We then have as an expression for the solids load: W = pQm¬] The work done per unit mass, P, on the mixture to move it along the pipe the

Jan 1, 1945

By Guosheng Zhan

The Homestake Mine included the open cut and underground workings up to 2,484 m (8,150 ft) deep. The mine ceased mining activities in December 2001 after 125 years of operation. Pumping from the underground workings continued for 18 more months until June 10, 2003. Homestake had completed a comprehensive hydrologic evaluation on post-closure underground flooding prior to the shutdown of pumping. After closure, the mine was selected by the National Science Foundation as the site for the Deep Underground Science and Engineering Laboratory (DUSEL). An interim laboratory facility will be set at the 1478 m (4850 ft) Level. The final facility will be set at the 2256 m (7400 ft) Level or deeper. Pumping from the underground workings began on June 5, 2008 by the South Dakota Science and Technology Authority (SDSTA). During the period of June 2008 through March 2009, the pumping rate increased from 41 L/s (650 gpm) to 81 L/s (1,290 gpm) with an average rate of 60 L/s (950 gpm). As a result, the water level had been lowered more than 95 m (310 ft). It was found that the observed rebound and later decline to date is in close agreement with the model projections. Because of this, the authors of this paper reject claims of a lack of sufficient measurement devices in place and a lack of understanding of the hydrologic system at the Homestake Mine prior to shutdown of pumping in 2003 (Davis et al, 2009). Furthermore, it is believed that this validated model can serve as a useful tool for the SDSTA to manage future underground dewatering for the DUSEL operation. This paper evaluates the hydrologic conditions in the Homestake Mine area and projects the post-closure groundwater rebound and decline due to DUSEL pumping.

Jan 4, 2010

By W. E. Lee

Penetration and dissolution mechanisms are reviewed for predominantly single-phase oxide, two phase oxide and oxide-carbon composite refractories by liquid silicate slags. Theoretical models of these processes, as well as static (sessile drop, dipping and crucible) and dynamic (rotating finger and rotary slag) experimental tests, along with their practical limitations are considered. Direct (congruent or homogeneous) attack is controlled by the reaction rate at the slag-refractory interface or the rate of diffusive transport of species to it through the slag leading to active corrosion. Indirect (incongruent or heterogeneous) attack is controlled by diffusive transport through the slag or through a new solid phase, which forms at the original slag-refractory interface. This may lead to passive corrosion. Examples of direct and indirect attack in a range of refractory/slag systems are described highlighting the critical influence of the composition and hence viscosity of the local liquid slag adjacent the solid refractory. Penetration and corrosion can be controlled either through the local slag composition via the refractory or the bulk slag or by microstructural control of the refractory by e.g. internal generation of dense layers or external deposition/generation of passive coatings, so-called in situ refractories.

Jan 1, 2004

By Patrick J. Hannigan

A comparison of 40 statically and dynamically load tested piles was performed to evaluate the variation between testing methods. This evaluation indicated an agreement of 10% to 15% could be obtained when the available static resistance is fully mobilized by both methods and when time dependent capacity changes are minimized. Dynamic pile testing recommendations for minimum induced impact force, ram weight to pile weight ratio, and considerations for time dependent capacity changes are presented.

Jan 1, 1987

By P Goblet, S A. Séguret, E Cordier, A Galli

In salar-type deposits, lithium grades can be measured by bailers introduced at different elevations in the drill holes at locations where, later, the brine will be pumped out to recover the metal. When they are duplicated, these a priori rudimentary measurements may show large differences without indicating whether this is due to the procedure itself or to some physical causes linked to the dynamic behaviour of the salar (seasonality, rainfall, underground flow etc). In the reservoir presented in the paper, the problem is complicated by the double lack of stationarity of the lithium grades: the grades not only increase importantly with the depth but at the same time their fluctuations decrease, making it necessary to use non-stationary geostatistical techniques to simulate them. Reservoir simulation is developed in two steps: first, geostatistical simulations of the lithium grade create possible realisations of the reservoir; then, each geostatistical simulation is followed by a pumping simulation that calculates, at each drill hole of a given extraction scenario, the lithium produced each year over 40 years. For the set of 100 simulations produced in this way, a scenario reduction is made to choose the five most representative simulations where more complex calculations will be carried out (for example, changing hydrogeological parameters such as porosity and permeability, or the elevation of the filters, the number and the location of the pumping and the reinjection drill holes etc). The paper first presents an analysis of the geostatistical model sensitivity to the grade uncertainty when measured by bailer. Among the 100 measurements at our disposal covering the future production domain, 50 per cent are duplicates measured at the same place but at different times. They are randomly sampled to produce five subsets of 75 values, which will constitute the future 3D conditioning points for the geostatistical analysis and the simulations. For each subset, trend, standard deviation and variogram models are fitted, leading to five sets of 20 geostatistical simulations of grades. In this way, the reliability of the different parameters involved in the procedure is evaluated, as well as its impact on the pumping results. Then, a reconciliation study is conducted between the non-stationary nugget effects encountered in the 3D variograms of the lithium grade and a pure statistical analysis of 25 duplicates measured at different locations and/or different times. The result is that if the nugget effect reaches 20 per cent, which is the case of three conditioning subsets out of five, the reconciliation is good and the nugget effect of the variogram model represents the bailer uncertainty, ie the measurement error. Since the geostatistical simulation incorporates a nugget effect, it handles the bailer uncertainty and the impact on the produced metal is included in the evaluation. The final conclusion is that, after the hydrogeological pumping simulations, the initial differences between the five geostatistical models do not influence the final results: lithium grades measured by bailers can lead to a robust evaluation of the extractable resources, a conclusion that conflicts with conventional wisdom.CITATION:Séguret, S A, Goblet, P, Cordier, E and Galli, A, 2017. Bailer uncertainty evaluation in a lithium salar deposit, in Proceedings Eighth World Conference on Sampling and Blending , pp 395–402 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By Tom Treleaven

The success of any blast is dependent on the accuracy of the blast design and the field controls used on the shot. This point cannot be over emphasized because when the geometry is not measured and/or field controls are not properly implemented, it can often have devastating effects on the final blast results. Thus drilling accuracy, tiont row burden control, explosive confinement, energy distribution, explosive selection, timing and structural geology become critical parameters in the design of successful blasts. Understanding the actual dynamics of the blast and why certain results were achieved is the fast step in optimizing a blast design. As stated by R. Frank Chiappetta, “Of all the blast monitoring systems available today, high speed videography still remains as the best stand alone instrumentation system for blast diagnostics.” Over the past years I have analyzed hundreds of blasts using either high-speed lowcams (500 frames per second) cameras or standard digital (30 frames per second). These analyses confirm, without a doubt, that shot geometry, in combination with structural geology, is the most important factor in overall blast results. Drilling accuracy is the single most important component in efficient blasting. Having equal mass associated with an equivalent energy is the basis in which blasts are designed. When drilling inaccuracy or borehole deviation occurs, not only is the energy unevenly distributed, but the basis of all design calculations becomes invalid. Over the past several years, quarry and mining operations have employed the use of boretracking measuring devices to measure borehole deviation and hole toe locations. In cases of operations using top hammer drills, it was discovered that the amount of deviation was significant and far greater than they believed they had. Set up error, improper drilling techniques and bit wear were some of the contributing factors. Through video anaysis of boretracked blasts, blast results were directly linked to drilling ..accuracy. The comparisons here took into account blast design, energy factors and hole sequencing.

Jan 1, 2000

By T. Tafi

The Ute-Ulay project area is located approximately five miles west of Lake City, in Hinsdale County, Colorado. The Ute-Ulay Mine site is approximately one mile east of the project area. The elevation of the site is 9,200 feet and is adjacent to Henson Creek (Figure 1) [ ] The Ute-Ulay Mining complex is an inactive gold, silver, and zinc mining/milling operation, that operated from the 1880?s until the 1970?s, and sporadically into the late 1990?s. The mining complex comprises several levels of underground mines, a flotation mill, and an abandoned town site. Due to the narrow, steep sided canyon, milled tailings were pumped upstream one mile in a sluice box to a wider section of the canyon, and deposited into five tailings impoundments and one berm located between the impoundments. An estimated 9,000 cubic yards (yd3) of fine grained tailings were located in the impoundments adjacent to the creek. The Ute-Ulay project area consisted of the mill tailings impoundments located upstream from the historic mine.

Jan 1, 2011

By Mathew Cook

Technological developments in Distributed Fiber Optic Sensing and improvements in both hardware and software have allowed fiber optic sensing systems to be deployed in a variety of applications such as pipeline leak detection, perimeter security monitoring, and power cable real time thermal rating. Recently, there has been a larger adoption of monitoring of conveyor belts in so called Conveyor Health Monitiring Systems (CHMS) within the mining and bulk material handling industries to identify idler conditions on an ongoing basis. Several attempts have been made to institute sensing capabilities into conveyors with mixed success. However, recent breakthroughs in modeling, installation methodology, and technology, as well as increased understanding of application and industry requirements, have allowed the technology to be deployed at both experimental and commercial scales. This paper summarizes the application conditions and requirements, design considerations, methodologies used, the outcomes achieved, the logistical and field considerations, and the future possibilities of the use of this technology in CHMS using Distributed Fiber Optic Sensing.

Feb 22, 2026

By Gray Mullins

The construction of drilled shafts as well as other cast in place foundation alternatives relies heavily on good practices from the contractor, engineer, and inspector in order to produce a quality foundation element. As many of the installation methods involve blind concreting processes, it is difficult to be certain of an intact concrete mass of the intended dimensions. A new integrity test has been developed that provides additional insight into the integrity of drilled shaft concrete. By measuring the temperature throughout the shaft via standard access tubes, the measured temperature profile can be compared to the normal signature associated with a shaft of the specified size. This article provides an overview of the new method development, test procedure, analysis, and results.

Jan 1, 2010

By C. K. Arthur, N. K. Dumakor-Dupey, B. A. Kansake, F. A. Kaba

Autonomous mining systems have been deployed for improving mine performance. However, there is a fear of job losses. Adoption of autonomous systems is envisaged to be resisted by major stakeholders in Ghana due to high illiteracy and unemployment. This paper assesses the preparedness of Ghanaian mining stakeholders for the adoption of autonomous systems to Ghana’s surface mining industry. Closed and open-ended questionnaires were administered to mine stakeholders. Respondents generally expressed unwillingness to accept autonomous systems into Ghanaian mines due to fear of increased unemployment. We propose setting up a mining education fund (MEF) for training stakeholders in autonomous mining systems. INTRODUCTION Mining contributes enormously to the sustenance and growth of global and national economies through the provision of raw materials, employment, revenues and foreign exchange (Carvalho, 2017; ICMM, 2017; Cordes et al., 2016; Solomon et al., 2008; Walser, 2000). The International Council on Mining & Metals (ICMM, 2014) reported that globally, formal mining employs approximately 2.5 million people. Many resource-rich countries, especially the developing ones, have experienced strong socioeconomic growth propelled by the discovery and extraction of valuable minerals. McMahon and Moreira (2014) reported a positive linkage between mining and sustainable development. They revealed that averagely, resource-rich countries have improved Human Development Index (HDI) scores than countries without mineral resources. Mining remains a crucial industry in Ghana’s economic development (Bermúdez-Lugo, 2016; Amponsah-Tawiah and Dartey-Baah, 2011). Over the years, mining has contributed significantly to the gross domestic product (GDP) of Ghana. In 2013, mining accounted for 1.7% of Ghana’s GDP and about 19% of government revenues (Ghana Chamber of Mines, 2015). The Ghana Chamber of Mines also reported that mining contributed over 50% of foreign direct investment, 37% of export earnings and 1.1% of direct employment revenues. The US Department of Energy (DOE) estimates that 65% of coal, 92% of metals and 96% of industrial minerals are mined using surface mining methods (DOE, 2007). Generally, mining is a capital-intensive venture, with high operating cost, hence there is a constant demand for mining companies to seek efficient cost reduction and production maximization strategies. One strategy being pursued keenly is automation, which is noted to improve productivity, and worker safety at minimal cost. According to Parreira (2013), mining automation involves direct teleoperation whereby workers remotely control the mining process using computers from a control room.

Jan 1, 2019