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By G. L. Gates

The Bureau of Mines in cooperation with the City of Long Beach) Calif.) studied some physical factors effecting the deposition of barium sulfate scale in oil-well rotary pumps. Freshly formed barium sulfate was obtained by combining sulfate in sea water with barium from a barium chloride solution. The rate of scale deposition was measured quantitatively by weighing the pump impellers and diffusers. The results of the tests showed that coating the impellers with plastic significantly reduced the rate of barium sulfate deposition.

Jan 1, 1967

By H. Sevim

Mining engineering curriculum has been discussed and debated in meetings and conferences, and opinions and recommendations have been published in the literature since the inception of the official degree-granting mining engineering program at the Columbia School of Mines in 1864. The deliberations have usually been heightened during the periods of boom-and-bust cycles experienced by the mining industry since the 1920s. Currently, the global natural resource industry is undergoing another cycle, called ?super cycle,? characterized by elevated prices for metals and other natural resources. The super cycle is occurring as a result of the large deficit between the global supply and demand for natural resources, due, in large part, to economic expansion and development in countries such as China, India and Brazil. Because of this cycle, there is a shortage of mining engineers in the U.S., Canada and Australia.

Jan 1, 2012

By T. Crivits

"Freeze lining technologies are used industrially to protect the integrity of furnace walls in pyrometallurgical reactors. This is achieved by solidifying a protective layer of slag on the inner furnace wall. In the current study, laboratory experiments have been undertaken to produce freeze linings in ‘Cu2232 O’-‘FeO’-MgO-SiO slags under controlled conditions using an air-cooled probe technique. Whilst most thermal models of freeze linings have assumed the freeze lining-bath interface to be at the liquidus temperature of the bulk bath and to be planar in geometry, the current research has demonstrated that this is not the case. Experiments undertaken using this high-temperature slag system have shown the interface at steady state conditions to be at subliquidus temperatures. The solid phases present at the interface do not include the primary phase and the interface is non-planar. The study indicates the need for improved understanding of the fundamentals of freeze lining formation, and of the influence of system properties and process parameters"

Jan 1, 2015

By Adil Jamali, Taufiq Hidayat, Imam Santoso, Zulfiadi Zulhan, Indah Suryani

The study investigates the segregation of iron and titanium in iron sand concentrate pellets reduced using coal under an isothermal-temperature gradient profile. The results show that metallic iron migrates to the surface while titanium and other oxides move to the center of the pellet. The study provides insights into the reduction process and the behavior of iron and titanium during heat treatment.

Mar 2, 2022

By Leo H. Glanville

UNTIL 1934, natural ventilation was depended upon in the mine of the Climax Molybdenum Co. at Climax, Colorado. In that year a 7-ft. axial-flow, low-pressure fan was installed as an exhausting unit. In 1936, three more fans of the same kind were installed, all as exhausting units, with a total capacity of 115,000 cu. ft. per min., and since then fans and sprays have been added as required. ATMOSPHERIC CONDITIONS Adverse atmospheric conditions at this elevation of 11,500 ft. account in part for difficulty in maintaining adequate ventilation throughout the mine. Barometric pressures range from 18.99 to 20.01 in. of mercury; the normal average being 19.70 in. Air weight is 0.051 lb. per cu. ft., and humidity is low-about a 30 per cent yearly average. This makes the use of mechanical humidity aids mandatory the year around to prevent excessive dryness of the mine workings and the resultant dust hazard. The yearly temperature range is from minus 24° to 95°F.; typically, in February 1942 the mean low was minus 3.25°F. and the mean high, 21.3°F. In August the mean low was 34.8°F.; the mean high, 64.26°F. Even during the summer month air filtration through the openings of the cave area is 32°F., owing to accumulation of snow in the glory hole and in part to glacial ice, which, combined with a rock temperature of 39°F. on the upper levels (elevation 11,945 ft.) and 42°F. on the main haulage level (elevation 11,490 ft.), makes it difficult to maintain a comfortable working temperature. The one advantage in ventilation of mine workings at the altitude of the Climax mine is that the power required to circulate a given air volume is 32 per cent less than would be required at sea level. PROBLEMS Since 1936 the mine management has studied intensively plans for providing adequate ventilation for the entire mine for its expected life. As the ore breaks into large pieces, much secondary blasting is necessary, therefore there is a considerable amount of smoke, gas and dust, which requires circulation of a large amount of air to remove the contamination. It is undesirable to use haulage drifts and manways as air passages because some would be passageways for contaminated air also, and there would be undesirable high velocities in some drifts. High velocities and large volumes of subfreezing air dry out the passageways through which they pass, and thus increase the amount of dust. Regulating doors in haulage drifts are difficult of upkeep and interfere with production. It was decided as a general principle to distribute fresh air to each working place at low velocities and low pressures and to remove contaminated air as soon as possible with a minimum of travel through working places.

Jan 1, 1943

By I. Vazaios, M. S. Diederichs, N. Vlachopoulos

"As rock engineering moves from continuum to more discontinuum approaches for modeling, the engineer must make decisions regarding the degree of complexity to simulate for a given application, balancing the level of structural detail within the rockmass model with computational cost and result accuracy. It is important to consider that model does not need to include all the existent fractures and that representative geometrical properties for a given rock mass volume can be defined including only major structural entities.Within this paper, the scale-dependency associated with the geometrical properties of fractured rock is investigated utilizing a Synthetic Rockmass (SRM) approach with a view to characterize such ground to aid in the design of geotechnical works. In terms of process, the concept of the Representative Elementary Volume (REV) is used to determine the equivalent structural properties of the rockmass in order to justify the applicability of far-field behaviour of such ground. A large size Discrete Fracture Network (DFN) measuring 60m x 60m x 60m has been generated using the structural data obtained from LiDAR scanning of the Brockville Tunnel, Ontario, Canada. A number of DFN samples of varying sizes are extracted from the master network, generated using MoFrac, in order to be used as a the basis of the structural analysis of the rockmass. Volumetric measurements for the density and the intensity of jointing are employed in order to approximate the geometrical REV for the networks.In this paper a complete methodology of how to generate a DFN from LiDAR data and its analysis in order to determine the size of a geometrical REV is demonstrated. For the specific rockmass found in the area of the Brockville Tunnel, a geometrical REV measuring 3.5m x 3.5m x 7m is established. The results are generally in good agreement with the in-situ observations."

Jan 1, 2015

By N. D. Jagger, I. M. Arbuthnot

Introduction Over the last five years, a large number of small- to medium-sized carbon-in-pulp treatment plants have been built in Australia, most designed to treat between 250,000 t/a and 1.5 Mt/a of ore. Because of the limited capital resources and tight cash-flow positions of these relatively small mining companies, the primary requirement was often to get a plant built and operating in a short period of time and at minimal capital cost. Therefore, since the inclusion of both pre-leach and tailings thickeners represents an obvious and significant capital cost, most of these plants were built without thickeners or even detailed, cost-benefit analyses on their inclusion. In some cases, the increasing use of High Rate Thickeners (HRTs) in the mineral processing industries has, however, resulted in a reassessment, because of their considerably lower cost. This reassessment was triggered primarily by the need to conserve water in arid mining areas where borefields are costly to install and water is limited. With the startup and operation of these installations, the resulting significant savings in cyanide consumption has been recognized, in many situations, as a primary justification for the installation of HRTs. Solution balances Degradation of cyanide occurs in the tailings water discharge to slime dams. The degree of degradation (cyanide loss) in the water recovered depends on a number of factors, but it is usually assumed to be about 90%. The most important mechanisms of CN loss are through HCN losses and oxidation by oxygen in the air, which also assists in the hydrolysis of CN. These mechanisms are supported by the large dam surface area and the long retention time of the tailings water in the dam. By thickening the CIP tailings at the plant and recovering as much tailings water as immediately possible, these losses are avoided. The retention time in an HRT is less than three hours, and the surface area is relatively small. Therefore, CN losses are negligible, which is not necessarily the case in conventionally-sized thickeners. Fig. I shows a block diagram of a 100-t/h gold plant without a thickener. In this example, 50% of the tailings water pumped to the tailings dam is recovered, and the CN concentration of the returned water is 10% of the tailings CN concentration of 150 ppm. Fig. 2 represents a plant with an HRT on tailings, thickening to 55% w/w solids. In this case, the thickened tailings are pumped to the dam, and 25% of the contained water is recovered. The recovered water and the mill make-up water are not sent directly to the mill; instead, they are added to the thickener feed and mixed with it prior to thickening. By doing this, the tailings are effectively washed, and the additional cyanide is recovered. Solution balances over these two circuits show cyanide recoveries of 5% and 65%, respectively. Thus, the thickener use increases cyanide recovery by 60%. Fig. 3 shows a two-stage HRT circuit in a countercurrent decantation (CCD) configuration. In this example, an additional 13% of cyanide is recovered through the use of the second-stage unit. This configuration can be justified when residual cyanide levels are high. Capital and operating costs - Case study 1 Illustrative cost figures are based on a CIP tailings thickener installed, in early 1988, as part of Dominion Mining's treatment plant at Paddy's Flat near Meekatharra. (All costs within this paper, unless stated otherwise, are in Australian dollars.) [Assumptions: Feed rate 150 t/h Ore moisture content 5% Leach density 40% solids Underflow density 55% solids Residual cyanide in tailings 150 g/m3 Flocculant dosage 15 g/t Consumable costs: Water $ 0.40/m3 Cyanide$ 2.00/kg Flocculant$ 4.50/kg Power$ 0.12/kWh Length of tailings pipeline2,500 m Capital costs of the major items involved in the thickener installation are given below: Thickener unit, 15-m diam $ 230,000 Water return pumps 12,000 Water recirculation pumps 28,000 Feed box 5,000 Flocculant make-up system 18,000 Flocculant storage tank and dosing pumps 11,000 Piping and valves 85,000 Electrics and instruments 47,000 Civil works 29,000 Installation 69,000 Total cost$ 534,000] Savings in capital costs that can be attributed to the

Jan 1, 1993

The long-term production schedule of open pit mines that maximises the net present value (NPV) of a given project relies on the proper design of phases (pushbacks) within the ultimate pit limits. There is not a phase design algorithm applied in the mining industry today that gives yearly production schedules that are optimum.The majority of phase design algorithms currently in use by the mining industry aims at grouping the blocks that fall within the ultimate pit limits into phases such that the average value of the blocks within each phase decrease as the size of the phases increase. Current phase design methods are typically based on a pit limit optimisation algorithm, originally introduced by Lerchs and Grossmann (LG) in 1965. To obtain multiple pit shells to be used for the phase design, the LG pit limit algorithm successively applied to economic block models by changing the price of the commodity or some other factor that properly parameterises the size of the pit. These modified economic block models are developed by predetermining the destination for each block in the model based on a breakeven analysis and highest economic outcome between alternative destinations. Phase designs determined by these existing algorithms may be suboptimal if a given project utilises multiple processing facilities for metal recovery and particularly if these facilities have restrictions (constraints) that require blending of attributes in the feed. To overcome these problems in obtaining phases that optimise overall production schedules, a new algorithm is developed at the Colorado School of Mines. The algorithm determines the phases of an open pit mine by taking into account multiple process options for the blocks and period by period blending requirements. This algorithm has been implemented within a software program (the software) and is being prototyped successfully in the design of improved pushbacks at a copper mine in South America and on a gold mine in Nevada. This presentation will demonstrate the shortcomings of the traditional phase design methodologies based on comparison of results coming from the new approach versus the traditional.

Dec 6, 2010

By William Gleason

"Sluggish prices for commodities and a glut of oversupply of many commodities did not stop the world’s largest mining equipment makers from showing off their latest and greatest machines and technology starting with the world’s largest equipment trade show, bauma 2016 in Munich, Germany.From April 8-14, an estimated 580,000 visitors from 200 countries visited the show that featured 3,423 exhibitors from 58 countries. While primarily a show for construction equipment, the mining footprint has grown over the years. Among the largest exhibits was that of Liebherr, which brought more than 100 products, had an outdoor exhibition stand, a stand for the components division and a career stand under the motto “Be Part of Something Bigger.” One of the biggest aspects of the exhibit was its mining products where the company staged the first-ever professional table tennis match in the 130 m3 (170 cu yd) dump body of the T 264 mining truck. The top German players Timo Boll and Patrick Franziska with their trainer Jörg Robkopf faced Austria’s singles world champion from 2003, Werner Schlager, and the reigning Team Europe champions Chen Weixing and Stefan Fegerl.“Bauma 2016 was an absolute success for Liebherr. With this year’s appearance we were once again able to show our innovative strength, performance and technological diversity to a very large and international audience from the industry. We conducted interesting discussions with numerous customers and partners from various countries and discussed new projects,” said Stefan Heissler, Member of the board of directors of the Liebherr Group.Likewise, Caterpillar, the world’s largest maker of construction and mining equipment had a large presence at the show with more than 60 products and 12,000 m2 (129,000 sq ft)of exhibit space. Among the iconic yellow mining equipment that was on display was the 6015B hydraulic mining shovel."

Jun 1, 2016

By S. Kelebek, U. Demir, T. C. H. Richards, E. Koss

"Laboratory-scale tests were carried out to investigate the response of a high-grade (~143 g/tonne) well-liberated gold ore to flotation using two collector types. Octyl and amyl xanthates, representing conventional types of collectors, were tested against an n-octanoic and hexanoic acids under moderate and high redox potential conditions. The latter was based upon a suggestion in literature that monocarboxylic acids are able to adsorb onto the surface of gold at redox potentials near 600 mV (SCE). This investigation did not demonstrate that these carboxylic acids could function as collectors for gold in real ores. In fact, they appear to have slightly depressed gold. However, the results demonstrated that octyl and amyl xanthates acted as highly effective gold collectors for this ore and that an increase in the collector dosage led to an increase in recovery for all mean particle sizes. In some tests, gold recoveries of around 95% were attainable in concentrates with grades around 6 kg/tonne, indicating that xanthate flotation remains a viable means of concentrating gold even at coarse grain sizes. Various aspects of the process such as grade-recovery performance and recovery by size are discussed.INTRODUCTION Three methods are used for the recovery of gold in mineral processing, either alone or in combination: gravity concentration, cyanidation and flotation. Gold occurs as a by-product in some ores together with platinum group metals and as the primary metal in other ores in the form of native gold and electrum, which are alloys of gold, silver and minor amounts of copper. Flotation is a viable method for recovering liberated gold where the gold is visible under an optical microscope (greater than 1 µm in size) and where the attachment of gold grains is not prevented due to bubble loading by competing minerals. It is generally known that grains finer than 150 µm in size will float with most collectors."

Jan 1, 2017

By Kentwell D. J, Cusak C. J, Harvey-Kelly F. E L, M D. Goodz

The Morning Star Gold Mine presents an example of a prolific operation with considerable historical records, mine workings, drill hole logs and countless plans from greater than 20 adjacent mines which have now been consolidated to form the current project. Implementation of 3D computer modelling in the planning stages has provided an additional tool to integrate and refine geological models as a guide to some exploration strategies. It has allowed for a re- interpretation of the host rock units, dimensional constraints, new exploration extensions and provisional resource modelling. The initial consideration of this project was to integrate 135 years of mining history into a single database. Each historic mining area had an unrelated mine grid, often established parallel to the long axis of their main stopes; some were based on magnetic or true north. Survey plans were in chain, imperial or metric units. Through the manipulation and rotation of old grids, previously unrelated mine workings began to form continuous mineralised zones, with unmined portions still remaining in between the existing development. Furthermore the current drillhole database of >433 holes showed numerous confirmatory intersections of new mineralisation. A series of tiered tables were created within a relational database to derive guidelines for assessing data to attribute a Mineral Prospectivity Value [MPV] based on key indicators of mineralisation. Indicator logging now forms the basis of geological descriptions.

Jan 1, 1997

By Tyler Luxner

Stillwater Mining Co. (SMC) is the largest primary producer of platinum and palladium in North America. SMC operates two underground mines in south central Montana, along the 45-km (28-mile) platinum/palladium J-M Reef in the Stillwater Complex. The Stillwater Mine exploits 9.6 km (6 miles) of deposit strike length and East Boulder Mine exploits 4 km (2.5 miles) of deposit length. The two mines are 20.4 km (12.7 miles) apart with SMC controlling a considerable length along the entire J-M Reef trace. Production at the Stillwater Mine began in 1986 and in 1988 the first tunnel boring machine (TBM) was commissioned, becoming an integral part of mine development at SMC. TBMs have been used for developing production footwall laterals (FWL) to access new mining blocks at the Stillwater Mine, as well as boring twin crosscut ac-cess tunnels at the East Boulder. Two TBM development expansion projects are currently in progress: the Graham Creek project at the East Boulder Mine, which will extend footwall lateral development 2,591 m (8,500 ft) to the west, and the Blitz project, where a TBM will be commissioned to extend east out of the Stillwater Mine to 3,658 m (12,000 ft) beyond the current production mining front. General strike and dip of the Reef is west to east and to the north, respectively.

Jan 1, 2012

By Ruiling DU, Jiazhi Zhang, Yong Zhao, Keng Wu

Reaction kinetics of metallurgical physical chemistry which was successfully applied in metallurgy (as ferrous metallurgy, non-ferrous metallurgy) became an important theoretical foundation for subject system of traditional metallurgy. Not only the research methods were very perfect, but also the independent structures and systems of it had been formed. One of the important tasks of metallurgical reaction engineering was the simulation of metallurgical process. And then, the mechanism of reaction process and the conversion time points of different control links should be obtained accurately. Therefore, the research methods and results of reaction kinetics in metallurgical physical chemistry were not very suitable for metallurgical reaction engineering. In order to provide the definite conditions of transmission, reaction kinetics parameters and the conversion time points of different control links for solving the transmission and reaction equations in metallurgical reaction engineering, a new method for researching kinetics mechanisms in metallurgical reaction engineering was proposed ,which was named stepwise attempt method. Then the comparison of results between the two methods and the further development of stepwise attempt method were discussed in this paper. As a new research method for reaction kinetics in metallurgical reaction engineering, stepwise attempt method could not only satisfy the development of metallurgical reaction engineering, but also provide necessary guarantees for establishing its independent subject system.

Jan 1, 2015

By François Robert, Steve Green, Tyson Birkett, Dan Richardson

Abstract -Detailed sampling using a portable diamond drill, and interlaboratory comparison of selected samples have been combined with statistical studies of truncated populations to establish the gold content of several outcrops in the Val d'Or area. The outcrops chosen are non-mineralized. The background gold concentration in the rocks of the region is 5 ppb or Jess. Gold is distributed relatively homogeneously within the outcrops studied except for slightly increased concentrations within quartz-tourmaline veinlets. Three samples, of which 20 grams of powder are analyzed for each, can adequately determine the gold concentration of a structurally simple outcrop in the Val d'Or area.

Jan 1, 1992

By Otto Braach, Wolfgang Frietzsche, Erdogan Gürkan

The Westerschelde Tunnel is located in the south of the Netherlands and links the two parts of the Provinces of Zeeland, Zeeuws-Vlaanderen and Zuid-Beveland, which are separated by the Westerschelde. Road traf?c has up till now taken the detour via Belgium or crossed by ferry. This unsatisfactory transport link signifies major economic disadvantages above all for Zeeuws-Vlaanderen. The first fixed link for road traffic between the communities of Terneuzen and Borsele are to be created by means of the two 6.6 km long tunnel tubes down to 65 m below sea level.

Jan 1, 2003

By PERCY E. BARBOUR

SECONDARY copper1 has &come more or less of a bugbear generally. What is its influence is often the subject of heated argument. The inedapable fact usually quoted is that since in 1929 primary production of copper in the United States was 1,026,348 tons and secondary production was 626,550 tons, there- fore, 0.6 lb. of secondary is produced for every pound of primary, and "How can you get around that?" The matter seems very plain, now that I have reached my conclusions, but they are the result of much laborious study and research. I shall show that' the production of secondary copper is not 0.6 per lb. of primary, in spite of statistical appearances; that the production of secondary is a function of primary production rather than of price; that secondary production has 'only a sub-. ordinate influence on the price of copper and then only in abnormal times; that there is not an accumulating reservoir of secondary from which the country is some- time to be drowned in copper, as is believed by some, or a reservoir from which we are going to draw much needed and more or less inexhaustible stores in the future, as is believed by others; and finally, that the United States always produces more copper than it uses and exports the surplus. These are all controversial points, and for that reason and to anticipate argument I have set forth my reasoning in considerable detail.

Jan 1, 1931

The JORC Code contains a mandatory classification system for Mineral Resource and Ore Reserve estimates for public reporting. The classification is a statement of the Competent PersonÆs confidence in the estimate and an attempt to convey that confidence to the various users of the resource estimate. In general the classification process is either largely subjective, relying on the Competent PersonÆs experience, or relies on statistical measures derived from the estimation process, or is derived from an estimated level of confidence in the outcome. However, the statistical measures may not necessarily relate to the overall confidence in the resource or reserve which must take account of the confidence in the geological inputs to the estimation.In addition to the traditional deterministic interpretation and estimations adopted by geologists, semi-quantitative techniques are now available to improve the estimation of confidence while continuing to involve the Competent PersonÆs experience in the process. Methods for quantifying confidence in mineral resource estimates for advanced projects and operating mines have also been developed.Classification is undertaken primarily as a communication tool between the estimators and the other stakeholders in the process. A brief examination of the stakeholdersÆ requirements is included.

Sep 26, 2011

By J. E. Armstrong

"IntroductionT'HE PRODUCTION of gravel and sand may be considered to be the major mining operation in the Lower Fraser Valley of British Columbia. At least thirty pits are in operation with a yearly production probably in excess of two million yards and a value in excess of one million five hundred thousand dollars before crushing, washing, and sizing. The value of the finished products would probably be double the above figure.Since 1949, the writer and his co-workers have been engaged in the geological investigation of the Lower Fraser valley and adjoining Coast mountains for the Geological Survey of Canada. The study, which is not yet completed, includes both bed-rock geology and geology of the unconsolidated sediments lying above bed-rock. In a rapidly expanding industrial area, .an advance knowledge of the geology may• greatly aid in future 'planning, and in many cases may .prevent not only the wastage of large sums of money, 'but possible disaster. Some of the present and future applications in which geological information in the Lower Fraser valley should be of value are as follows: the development of domestic, irrigation, industrial, and municipal groundwater supplies; drainage for industrial sites,, roadbeds, and farmlands; flood control along the Fraser river and smaller streams in the area; the location of dam sites, of major tunnels for water supply •and sewage, of road-beds, of foundation sites for large buildings .and other engineering undertakings in which geology may aid; the evaluation of oil and gas •possibilities; the production of building stone; and the production of sand and gravel, the subject with which this paper is primarily concerned."

Jan 1, 1953

By Q. H. Liao, B. G. Monaghan

The options for designing slopes in complex geological conditions are reviewed. It is shown that modelling, especially numerical modelling, allows a degree of analysis that cannot otherwise be achieved. Examples are presented of how numerical modelling can be used to investigate fundamental aspects of rock mass behaviour. It is demonstrated that an important consequence of change in scale is that the mechanisms of failure also change. The complex failure of the footwall slope at Aznalcollar, a large open-pit mine in southern Spain, is discussed. Numerical modelling with UDEC (universal distinct-element code) has allowed the realistic simulation of displacements measured in the field

Jun 18, 1905

By Z. Zhang, M. S. Gonzalez, M. de P. Bueno, P. Holopainen, S. Luukkanen

In an extremely competitive mining industry, onsite experience is a big advantage. Mining education at the universities is mostly focused on theoretical studies without a possibility of practical training in mining sites. Hence, experimental mines suitable for practical education are needed to provide a platform for systematic research and education in industrial scale and for training in real mining conditions. Yet, this kind of mine sites is rare worldwide. Thus, a new educational research project, namely MINETRAIN is introduced in this paper evaluating the transition of the Pyhäsalmi mine in Central Finland from an active base metal mine to a research, educational and training underground facility. The uniqueness of MINETRAIN compared to other test mine programs is that the existing state‐of‐art infrastructure in Pyhäsalmi enables research and training facilities among all disciplines related to the overall mine value chain. Though all the above sound interesting in the context of research and education purposes, in practice Pyhäsalmi will have to become an experimental mine that can be sustainable in the future. Accordingly, a prefeasibility study is being conducted and some preliminary results are presented in this paper. INTRODUCTION Despite being a field of studies that requires practical knowledge and experience, mining engineering education follows the same pattern in most mining universities around the world; consisting of theoretical courses and laboratory-based practical modules (Mischo, 2015). The possibility of having practical training in real mining conditions on actual mine sites or processing plants is limited to none, while the only contact that undergraduate students have with mining operations during their time of studies is usually through short-time internships or visits in the context of educational trips. Yet, real integration between universities and the mining industry may be scarce, which does not help students to observe the actual processes, and thus understand the whole picture of mining operations. Indirect practical training is also possible nowadays that technology is shifting from mechanical to digital, and the use of virtual reality -in the context of Mining 4.0- can provide additional experience to the learning process. However, virtual reality is limited to the visual impression and is also limited in actions (Binder et al, 2018). In any case, when there is not a proper pragmatic view of how the different disciplines link to each other and especially if there is a lack of communication, unfortunately the results could be seen at a later stage in the mines as a poor performance of professionals and workers. Another major challenge for educating specialists in the mining sector is that experimental mines suitable for practical education, where the skills and know-how could be developed and/or enhanced are rare worldwide. Such research and educational (R&E) facilities are practically former metal ore mines that have been transformed into experimental sites. Besides being limited in number, not all facilities of this kind serve both a research and educational purpose.

Jan 1, 2019