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By Lewis L. Oriard

INTRODUCTION In recent years, there has been a trend in the direction of larger drilling equipment and larger diameter blastholes. Although this change has improved the efficiencies and reduced the costs in many operations, it has increased the potential for damage to underground openings. In addition, in many instances one now finds more sophisticated delicate instruments, automated control facilities, and a large variety of structures in proximity to blasting activity. The combined effect of larger-scale blasting activity and its proximity to various features of interest is such that there is an increased need for a more refined analysis of blasting effects and their control. BLASTING EFFECTS ON ROCK SURFACES The Breakage Mechanism In order to develop techniques for controlled blasting, one must first understand the features of the mechanisms by which blasting causes rock breakage to occur. These features have not been easy to demonstrate, mostly due to the difficulty in making tests and observations at the high stress levels and short time durations involved. When an explosive charge is detonated, the material surrounding the charge is subjected to a nearly instantaneous, very high pressure [on the order of 1.4 to 13.8 GPa (0.2 to 2.0 X 106 psi), depending on the explosive]. If the charge is coupled to "average" rock, this pressure will pulverize the surrounding rock for a distance on the order of 1 to 3 charge radii in hard rock, and to a greater distance in softer rock (this is also dependent on the type of explosive). As the pressure wave passes into the rock, high tangential stresses cause radial cracks to appear, and the nearly discontinuous radial stress zones gen¬erated by the shock front may cause tangential cracks to appear. The extent of these cracks depends on the energy available in the explosive, how quickly the energy is transmitted to the rock, and the strength properties of the rock. The discontinuous shock front is quickly dis¬sipated, but the expanding gases generate a longer-acting pressure. A compressive pulse travels to the nearest face or internal rock boundary where it is reflected in tension. The tensile strengths of most rocks are roughly 40 to %o of their compressive strengths, so the rock may now fail in tension whereas it may have been able to support the diminished compressive phase without failure. The ten¬sile deflection typically produces a failure described as tensile slabbing or scabbing. Laboratory experiments and field experience have pretty well established that several mechanisms are involved. These include (1) the classical case of tensile parallel slabbing when the pressure pulse is reflected at a free surface; (2) failure under quasi-static compressive loading (the shape is normally irregular due to discontinuities in the rock); (3) radial cracking under the action of tangential stresses at the periphery of the expanding pressure pulse; (4) peripheral cracking at the discontinuous shock front which is quickly dissipated; and (5) additional mass shifting due to the venting of the explosive gases. The first three items have received much attention in the laboratory and the literature. The complex effects of gas venting are difficult to test in the laboratory because of the difficulty in reproducing the many weak planes and discontinuities typical of most field conditions, which play such a prominent role in determining the behavior of the rock mass subjected to blasting. Unfortunately, gas venting effects can be pro¬jected to significant distances under certain field conditions, and are sometimes difficult to control. It is not unusual for gas venting to be the overriding factor in determining the final geometric shape and physical condition of the finished excavation. Sources of Damage For the purposes of this discussion, damage includes not only the breaking and rupturing of rock beyond the desired limits of excavation but also an unwanted loosening, dislocation, and disturbance of the rock mass the integrity of which one wishes to preserve (such as mine pillars, underground openings, etc.). The sources of damage include, of course, all those physical features of the rock breakage mechanism. Each of these effects must be limited to the desired zone of breakage and excavation if the integrity of the remaining rock mass is to remain undiminished. The primary zone of rock breakage usually can be controlled in the normal process of field experimentation to determine proper charge sizes and location for primary excavation. However, it frequently happens that there is damage from sources which are more difficult to account for in the design process, which are often overlooked. These are (1) the overbreak due to poor drilling control, (2) dislocation of rock (mass shifting) due to venting of explosive gases, and (3) loosening or dislocation due to the influence of seismic waves (ground vibrations). CONTROL OF ROCK BREAKAGE Importance of Geometry In studying the rock mass and blasting design con¬siderations, it is important to keep in mind the geometric relationships among charge size, shape, and position, and the physical features of the rock mass to be preserved. The features of principal interest are the external shape and position of the rock mass relative to blasting, and the position and attitude of any weak planes in the rock mass. The Sequence of Blasting and Excavation Events Unfortunately, there are too many times when the task of preserving delicate rock is considered hopeless, and because of this attitude, no further effort is ex¬pended towards caution or control. In such cases there is often a failure to recognize the importance of the se¬quence of the procedures. Attention to this can greatly reduce unwanted effects at minimum cost. Perimeter Control The requirements for perimeter control are highly dependent on the special needs of each particular proj¬ect. The desirable degree of control is a highly variable

Jan 1, 1982

By J. Lovering, H. Wilhelm, P. Siewert

Introduction The Afton property is located 290 km (180 miles) by air east-northeast from Vancouver and 14 km (8.7 miles) west of Kamloops, a city of 60,000 people, in south central British Columbia, Canada. The mine is adjacent to the Trans-Canada Highway at an elevation of 670 m (2198 ft) above sea level. The ore body is a porphyry copper deposit that has undergone supergene alteration. The major economic minerals in the supergene zone are native copper and chalcocite with chalcopyrite and bornite in the hypergene areas. The grade is 1% with an overall copper distribution - 70% native, 25% chalcocite, and 5% chalcopyrite with bornite and covellite. The ore also contains important but variable amounts of gold and silver. The mill was designed to treat 6350 t/d (7000 stpd). Semiautogenous grinding was selected to minimize capital cost and because of the expected high clay content of the ore, which would have caused problems in a conventional crushing and screening plant. Test work indicated that a recovery of 87% was possible in a circuit incorporating both flotation and gravity separation. Flowsheet Run-of-mine ore is crushed in a 1.06 x 1.65-m (3.5 x 5.4-ft) Allis Chalmers gyratory crusher set at 228.6 mm (9 in.), closed side setting. The surge pocket, below the crusher, is emptied by a Hydrastroke feeder onto number one conveyor, which discharges onto a 180,000-t (198,416-st) coarse ore stockpile. Six Hydrastroke feeders on two conveyors withdraw the crushed material from the bottom of the pile. These two conveyors, in turn, discharge onto the belt feeding the semiautogenous mill. The live storage in the stockpile is approximately 22,000 t (24,250 st), sufficient for three days' mill feed. Primary grinding is accomplished in an 8.5-m (28-ft) diam by 3.7-m (12-ft) long Koppers (Hardinge Cascade) mill (Fig. 1) containing a 10% ball charge and driven by a 4000-kW dc variable speed motor. The mill dis¬charge is pumped by a 10 x 12 G.I.W. pump to a 1.22 x 4.88-m (4 x 16-ft) stationary screen sloped at 20°. Screen oversize returns to the semiautogenous mill (SAM), and the undersize flows by gravity to the ball mill discharge pump box. Secondary grinding is performed in a 5-m (16.4-ft) diam by 8.84-m (29-ft) Koppers overflow ball mill driven by a 3430-kW synchronous motor through an air clutch. The mill is in closed circuit with a Krebs Cyclopac containing 10 635-mm (25-in.) cyclones and the cyclone overflow, at 35% solids and 65% to 70% -200 mesh, is flotation feed. In order to limit the buildup of native copper, circulating in the secondary grinding circuit, a portion of the underflow from the cyclones is processed in a circuit containing screens, cyclones, and shaking tables to produce a finished metallic copper concentrate. Primary mill variable speed drive The overall waste to ore ratio at Afton was 4.5:1. The mining was to be done with only three shovels, which meant that it was highly unlikely that more than one of them would be in ore at any one time. The resulting inability to blend the mill feed made it impossible to prevent wide swings in the grade and grindability. The variable speed do drive motor installed on the semiautogenous mill was selected because of the extreme variability of the Afton ore body. This variability has persisted throughout the lifetime of the mine. There are times, however, when due to ore conditions, the mill is operated at full speed (78% of critical) for extended periods of several shifts duration. There are other times when the mill speed may be changed several times in a 12-hour shift due to changing ore conditions. When ore is processed that contains a fairly large proportion of fine native copper, the primary mill speed and, consequently, the tonnage may be reduced to improve the secondary grind and to maintain an acceptable grind and recovery. High clay ores require less mill speed and more dilute grinding densities. In the latter case, the slower primary mill speed also helps to minimize damage to the mill liners. Approximately 57% of the time the mill operates between 90% and 100% of full speed or between 71% and 78% of critical. The variable speed is also used for inching during mill relines.

Jan 1, 1987

By W. E. L. Minter

Discussion by R.W. Hutchinson Minter's excellent recent paper (1990) on the great paleoplacer gold deposits of the Witwatersrand and its subsequent discussion (Cheney, 1991) and reply (Minter, 1991) provide an informative, interesting and up-to-date review of the geology of these unique ores and some long enigmatic questions regarding their origin. Clearly summarized by Minter (1990) are the compelling sedimentologic and stratigraphic relationships indicating that the Central Rand Group conglomerates are ancient placers and documenting the fluvio-deltaic conditions of their transport and deposition. As Minter (1991) indicates, Cheney's discussion broadens Minter's 1990 coverage of the subject by considering higher and lower stratigraphic units and by adding detailed comments concerning recently published information about metamorphic and age relationships. Cheney's discussion and Minter's reply also readdress the old controversy of epigenetic hydrothermal vs. syndepositional placer origin for these deposits. This controversy has been resurrected by: • recent recognition of hydrothermally-altered granites (HAGS) in the older Archean basement rocks that lie unconformably beneath the Dominion Group rocks north and west of the Witwatersrand depository; • recognition of greenschist metamorphic assemblages in the Witwatersrand strata; and • the presence of interpreted epigenetic mineral assemblages and textures in some of the deposits. These relationships are briefly discussed by Cheney who suggests they may result from later, superimposed metamorphism and deformation and do not prove that the gold was introduced by epigenetic/metamorphogenic processes. In his paper, Minter (1990) cited the 1986 abstract of a subsequently published paper by Hutchinson and Viljoen (1988) who also considered many of these aspects and attempted to integrate them all in abroad genetic hypothesis. In addition, they emphasized the additional important problem of determining the source for the Witwatersrand gold. Although this issue was partially addressed by Reimer (1984), Mossman and Dyer (1985) and Reimer and Mossman (1990), the question has been generally under emphasized in earlier work and was, therefore, not stressed by Minter and Cheney. Reimer (1985) also emphasized the importance of weathering of the basin's Archean hinterland and its affects on Witwatersrand sedimentation. Since the epigenetic hydrothermal hypothesis was essentially discarded in favor of placerist origin, the gold source for these great deposits has generally been attributed to the weathering of lodes in greenstones or granites of the uplifted, older Archean hinterland to the north and west. However, Hutchinson & Viljoen (1988) present evidence that this explanation is inadequate, as well as data in support of an alternative explanation. They suggest that the gold and accompanying abundant, clearly detrital and auriferous pyrite in the conglomerate reefs were derived by erosional degradation of auriferous-pyritic exhalite previously deposited, by shallow marine discharge of hydrothermal fluids, along the fault-bound northwestern margins of the Witwatersrand depository. The auriferous-pyritic exhalites were generated by fluid-rock reactions in the subsurface, particularly in the highly-altered lavas of the Dominion Group but also in adjacent basement granitic rocks. The process envisaged is comparable to that observed in sea floor hydrothermal systems today, although, in this case, it occurred at the margin of a rapidly subsiding, shallow marine or continental basin. Important evidence for this explanation is the anomalous gold content of the ferruginous, shaley strata occurring throughout the Witwatersrand Supergroup and including the distinctive contorted bed that is an iron formation. These layers have been interpreted (Hutchinson and Viljoen, 1988 and 1990) to represent the more distal, finer-grained, mixed chemical/clastic sedimentary strata. Thus, they are the basinal equivalents of the proximal, auriferous-pyritic exhalites that were deposited along the basin's margin and subsequently reworked by sedimentary processes. This gold source explanation resolves many of the enigmatic questions regarding Witwatersrand geology and the reawakened controversy between proponents of hydrothermal and placerist genetic hypotheses. The theory invokes a hydrothermal origin for the gold in the basin-margin pyritic exhalites, thus explaining the visible hydrothermal characteristics (Figs. 1 and 2). It also provides compelling evidence for placer origin by encompassing all of the fluvio-deltaic processes of clastic sedimentary transport and deposition that are clearly summarized by Minter. These characteristics result from the degradational reworking of the pyritic exhalites along the depository's margins. Fluvial reworking, transport and deltaic redeposition of this detritus, along with additional detritus carried into the depository from the hinterland, then formed the conglomerate reefs. Other enigmatic aspects of Witwatersrand geology include: • the rarity of magnetite and ilmenite (the two most common heavy minerals in placer deposits) and the contrary abundance of leucoxene in the conglomerate reefs; • the presence of differing varieties of pyrite (authigenic, allogenic, compact and porous) all of which are auriferous; and • the presence of well-rounded pebbles, even cobbles, of pyrite that, in view of the brittleness of the mineral, could not have been transported great distances from a hinterland.

Jan 1, 1993

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

5.1 TECHNOLOGY FOR INJECTING GROUT INTO BOREHOLES The grouting of fractured rock aquifers is carried out in separate stopes' whose size is determined by the hydro¬stratigraphy and the hydrogeologic properties of the rocks. The number of grouting stopes required depends on the number of intervals in the hydrogeological section. Each grouting stope can include one or several aquifers. Com¬bining several water-producing strata into one stope is pos¬sible only when the values of their permeability coefficients are very similar. In order to achieve efficiency, as many aquifers as possible are included in each stope. In practice, the length of a grouting stope in a borehole usually ranges within 10 to 40 m, but occasionally up to 100 m or more. When conducting grouting operations through boreholes drilled from the face of underground workings, the stope's size can be determined by the thick¬ness of each fractured zone (fault) or by the length of the hole if the entire rock is fractured. Injecting grout into separate intervals of fractured rock can be done from the bottom to the top with ascending stopes or from the top to the bottom with descending stopes. Injection from bottom to top is always preferred. Injection also can be accomplished in separate fractured permeable horizons in the section. The STG integrated grouting method is based on the use of packer and injection equipment and technological schemes that facilitate injection in highly isolated fractured aquifers. For this purpose STG has developed a standard series of packers. 5.1.1 INJECTING GROUT THROUGH BOREHOLES DRILLED FROM THE SURFACE Grout can be injected according to one of four techno¬logical schemes depending on the depth of the fractured aquifers, their hydrogeologic properties and the availability of grout injection equipment if the area is remote (Fig. 43). In the first and most preferred technological scheme (1), the grout is injected from the bottom up with ascending stopes. When this scheme is employed, grouting holes are drilled to the entire projected depth based on the field investigations injection tubes is installed above the top of the first hydrostratigraphic unit to be grouted. This scheme al¬lows the grout to be injected only into the permeable stra¬tum or strata isolated by the packer near the bottom of the hole; it precludes the need to clean or redrill the hole before injecting grout into the next overlying stope. This ascending stope scheme is preferred because of its simplicity and ef¬ficiency. It is used in most of the grouting of saturated fractured rock for the purpose of isolating aquifers while excavating deep mine shafts in the USSR. The second technological scheme (II in Fig. 43) injects the grout in sequence from top to bottom with descending stopes. Injection under this scheme is accomplished through the hermetically-sealed mouth of the hole. It is most advis¬able to grout with descending stopes in saturated fractured rock wherein the aquifers constitute thick networks of frac¬tures. Under these conditions it is practically impossible to drill a grout hole to the entire depth of the underground workings due to the large production rate of ground water. In conducting isolation operations with descending stopes, additional work (relative to ascending stopes) is necessary to drill out the grout in order to create and inject the next stope. As a result, the total amount of drilling required to drill out grout can even exceed the volume of drilling of original rock. Consequently it is advisable to use this inefficient scheme only for injecting grout into holes that contain prolific frac¬tured aquifers at depths less than 150 to 200 m. The third and fourth technological schemes (III and IV in Fig. 43) are used in the selective grouting of separate aquifers. For example, these methods would be applied when a grout hole is drilled to its entire depth, wherein grouting of the upper aquifers must be done first so that the sinking of the first shaft can begin prior to deeper grouting. The third technological scheme is used for the selective injection of grout into an aquifer with a thickness of up to 5 m. For this work the DAU-1-2 removable dual packer is installed. It is lowered into the borehole on an injection pipeline. For aquifers with thicknesses of more then 5 m, the selective injection of the grout is carried out according to the fourth technological scheme; a combination of the DAU-2 hydrodynamic plug and the DAU-1 packing device is used. In order to implement this injection scheme, the DAU-2 hydro-dynamic plug is installed at the bottom of the fractured aquifer and the DAU-1 packer is installed at the top of the aquifer.

Jan 1, 1993

By Rosa H. Harris

The world's largest gold dump leach operation, Minera Yanacocha S.A., has installed community assistance programs in Peru concerned with health, education, electrification, agricultural assistance and road improvements. The programs are conducted in a joint venture fashion in association with various organizations.

Jan 1, 1998

By Lawson. Heather

"The purpose of this study is to explore how the geochemical and petrographic components of coal may impact its physical properties, as well as how these components correlate with a history of reportable dynamic failure in coal mines. Dynamic failure events, also termed bumps, bounces, or bursts, are the explosive failures of rock in a mining environment. These events occur suddenly and often with no warning, resulting in worker injury up to and including fatality in greater than 60% of MSHA reportable cases. A database of variables was compiled using publicly available datasets, which included compositional geographic, strength, and Hardgrove Grindability Index (HGI) data (ACARP, 1998). Results indicate that bumping coals are less mature, lower in carbon, higher in oxygen, softer, and less well cleated than coals that do not bump. High liptinite content was found to correlate with higher average UCS values. However, no clear, direct correlation between UCS and dynamic failure status was observed. The findings of this study establish that differences exist between coals that have experienced reportable dynamic failure accidents versus those that have not. These differences are inherent to the coal itself, independent of mining-induced risk factors. Results further illuminate how compositional attributes of coal influence physical properties and begin to clarify potential links between geochemistry and dynamic failure status. Only through the better understanding of risk can more effective mitigating strategies be enacted.INTRODUCTIONDynamic failure events, also termed bumps, bounces, or bursts, are the sudden failures of rock in a mining or quarrying environment. Failure occurs when the rock’s critical bearing capacity has been exceeded and the rock fails energetically through the outward expulsion of rubblized material (Peng, 2008). These events occur suddenly and often with no warning, resulting in worker injury up to and including fatality in greater than 60% of reported cases through the Mine Safety and Health Administration (MSHA). Much research has been devoted to the prevention of these events. The effects of overburden depth and stiffness, mine design, mining practices, and in situ stresses are well documented (Agapito and Goodrich, 2000; Brauner, 1994; Campoli and Kertis, 1987; Holland and Thomas, 1954; Iannacchione and Zelanko, 1995; Lawson et al., 2016b; Mark, 2016; Newman, 2002; Peng, 2008; Whyatt, 2008). Despite these significant advancements in coal mine ground control, however, events continue to occur. Proactive risk mitigation remains an important research area. This is particularly true in underground coal settings, where the rockmass is layered, lithologically diverse, and does not exhibit consistent material properties across deposits.Many conditions have been associated with the occurrence of dynamic failure phenomena, including:"

Jan 1, 2019

By J. Herhold, S. Plaum

"Current and future operational processes in the mining and underground construction sector are under the influence of the prevalent competition and the acute necessity of optimization approaches concerning corporate and project processes. Besides general construction procedures, especially underground construction procedures and procedures in the field of deep shaft construction are high-profile. Those processes describe high technical challenges as well as increased requirements to occupational safety and health. Therefor an occupational safety management, which should be adjustable to the company itself and adaptable to all kinds of projects, shall be created. The creation of such an occupational safety management would lead to an economically optimized and safe (within the scope of OSH law) implementation of internal activities for the performance of construction services underground and in the field of deep shaft construction. With this approach the satisfaction of the high safety request as well as the demand for process optimization, as part of keeping the own competitiveness, will be satisfied successfully at once. PROJECT BACKGROUND Introduction The health and safety protection as well as occupational safety look back on a long historical development, which had brought more important sectors and developments during the last years, not only in Germany but also on the European and international level. The accident insurance protected by law has already been reported in 1884, which is astonishing early. The launch time point is strongly connected with the on-going economic development in the end of the 18th century, but above all this increased in the 19th century. The progress of occupational safety and especially the interest in the integration and the concentrated consideration of occupational safety and the health and safety protection in the company process is strongly visible during the 19th century and directly connected to the industrial revolution. [9] The industrial revolution and with that the development of work places lead to changing demands in the field occupational safety and health and safety protection. Insufficient security precautions caused problems in the industrial revolution, such as personnel and material losses as well as serious injuries and their consequences. These and the fundamental economic endeavours of the entrepreneurs lead therefore to the important step to develop an occupational safety system. [9] The prevailing basic ideas seize the insufficient security precaution of work places as well as the economical and profitable ideas of the entrepreneurs and bring them together. From this the necessary entrepreneur’s task comes into being and that is bringing the work conditions to the way that life and health of employees are not in danger. [9]"

Jan 1, 2017

By Sadatoshi Ohmori, Karen Shinoda, Nobuharu Isago

Method of tunnel excavation is generally classified such as full-face excavation, bench-cut excavation, diaphragm with center pillar excavation, center drift advancing excavation and so on. These excavation methods should be selected properly in terms of ground condition, the area of tunnel section, the arrangement of construction. Center drift advancing excavation method in mountain tunnel construction, is often applied under quite weak ground condition, prior to the main tunnel widening excavation. Recently, the advancing drift is for the excavation of extreme larger section in main tunnel to confirm the stability of face. In addition, we make advancing deformation occur in drift when the construction was done in poor ground like squeezing and swelling ground, and the force occurred in main tunnel was expected to reduce by the effect of ‘dodging’ for support structure. Mechanical relation between drift and main tunnel has been mainly grasped on past experiences in Japan. However, the characteristic of tunnel enlargement was partially discussed by using the three dimensional analysis and the result was grasped by past experiences on the basis of each tunnel construction. It is considered that the natural ground stress is normally released as the advance displacement, but it was released after the main tunnel widening, causing a displacement and deformation increase. Up to now, the construction practice has shown the reduced displacement during the main tunnel widening excavation by the drift, and the numerical analysis agreed with above observation. However, there is no report found that the displacement and deformation at the time of widening excavation of the main tunnel become larger in case of a drift application. In Kyusyu area, western part of Japan, the northern part of Yoshinomoto tunnel was applied an advancing bench-cutting method for a center drift for the purpose of obtaining the geological information for support selection of the main tunnel in widening excavation. The shape of center drift was a perfect circle with a diameter of 4.0 m or 6.0 m and the same temporary support regardless of the geological situation. As a result, the displacement and deformation of the main tunnel after widening excavation were larger than the section where the center drift was not applied. Furthermore, the displacement and deformation of the main tunnel after its widening excavation was considered larger in the section where the cross-section of the center drift with 4.0 m diameter was applied than one with 6.0 m. Therefore, the center drift with a high rigidity support structure had the effect of suppressing the advancing deformation prior to the main tunnel widening excavation.

Jun 13, 2021

By T. M. “Mike” Gurr

"U. S. marketable phosphate rock production for the 2018 crop year (July 1, 2017 to June 30, 2018) showed an increase to 27.7 Mt (30.5 million st) from the 2017 crop year production of 27.1 Mt (29.8 million st), as reported to the U. S. Department of the Interior by the mining companies operating in the United States during crop year 2018 (Table 1). Domestic consumption decreased in the 2018 crop year to 23.9 Mt (26.3 million st), as compared to the 27.2 Mt (30 million st) consumed in crop year 2017. U.S. imports of phosphate rock increased to 3 Mt (3.3 million st) for the 2018 crop year from the 1.81 Mt (1.9 million st) for crop year 2017. Imports of phosphate rock were from Peru and Morocco. There was an increase of 2.85 Mt (3.1 million st) in producers’ stocks from 7.45 Mt (8.2 million st) in crop year 2017 to 10.3 Mt (11.3 million st) in crop year 2018 (Table I).For calendar year 2018, Florida and North Carolina phosphate rock production accounted for 75 percent of the total U.S. production, with the balance being produced in Idaho and Utah. The Florida and North Carolina percent of U.S. production is declining. U.S. production of 25.7 Mt (28.3 million st) in calendar year 2018 was produced by five companies at 10 mines in four states and accounted for approximately 9.6 percent of the world production of 267.7 Mt (295 million st) for calendar year 2018 (Table 2).In calendar year 2018, U.S. phosphate production decreased by 2.2 Mt (2.4 million st), while companies increased stocks of phosphate rock by 1.7 Mt (1.9 million st). The domestic production for calendar year 2018 of 25.7 Mt (28.3 million st) represents 78.4 percent of the domestic production capacity of 32.8 Mt (36.1 million st).At present, China reportedly produces more than 50 percent of the world’s phosphate rock. As a result of planned expansions, Morocco production has increased from 27 Mt (30 million st) in 2017 to 33 Mt/a (36 million stpy), which is approximately 12.3 percent of the world’s production, moving ahead of the U.S. production. In calendar year 2018, the U.S. phosphate rock production was 25.7 Mt (28.3 million st), which is less than 10 percent of the world’s production of phosphate rock (Fig. 1)."

Aug 1, 2019

By H. -S. Yoon, C. -J. Kim, K. W. Chung

"With the final goal of ultrasonic application to ion exchange in a resin-in-pulp (RIP) system, the feasibility of an indirect ultrasonic application with a submersed ultrasonic probe was investigated first for the enhancement of ion exchange in the following resin-in-solution (RIS) system: Lewatit MonoPlus MP 600XL in an artificial solution with 106 ppm uranium. Ultrasonic radiation enhanced the ion exchange efficiency without mechanical degradation of resin. For 3.6 g/L of resin, the enhancement was 79 percent after 10 min of ion exchange and it gradually decreased to 11 percent at 60 min. In addition, a preliminary test of the mechanical stability of the resin in the RIP system showed little mechanical fracture of the resin due to ultrasound, which indicates that indirect ultrasonic application by a submersed ultrasonic probe is applicable for RIP systems. IntroductionIn uranium ore processing, the ion exchange process has been used on a commercial scale for preconcentration and purification (Gonzalez-Luque and Streat, 1984). The ion exchange process is divided into two processes: (1) a resin-in-column process with the uranium-bearing solutions passed through columns containing resins and (2) a resin-in-pulp (RIP) process with the leached ore pulp exposed to resins. The latter does not require the separation process of leach liquor and ore solids, which reduces process costs. This process, which has been developed commercially since 1955, has a certain drawback. The adsorption rate and efficiency in the RIP process are relatively low, so numerous contacts between resins and leached ore pulp are needed, which adds to the cost (Merritt, 1971). It is necessary to study methods to reduce process costs.Ultrasound can intensify mass transfer, increase the rate of chemical reactions, and change reaction pathways (Thompson and Doraiswamy, 1999; Luque de Castro and Priego-Capotem, 2007; Adewuyi, 2001; Luque-Garcia and Luque de Castro, 2003). It has great potential in chemistry and is especially considered to be a promising method for enhancing mass-transfer-limited processes, such as ion exchange. Enterzari and Tahmasbi (2009) reported the effect of ultrasound on water softening by ion exchange. Band et al. (1997) proposed a modulated ultrasound application to water desalination with ion-exchange hollow fibers. These reports revealed that ion exchange kinetics could be enhanced by the application of ultrasound. As for hydrometallurgy, Feng and Aldrich (2000) reported that the elution of an ion exchange resin was enhanced by the use of ultrasonication."

Jan 1, 2015

By John K. (Original by) Berry

The handling and disposal of mine water is a much larger problem than is apparent at first glance. Many more tons of water are removed from underground coal mines in the United States each year than tons of coal. In the old anthracite region of Pennsylvania the average was around 32.7 t (36 st) of water for each ton of coal produced, while bituminous mines average about 4.5 or 5.4 t (5 or 6 st). Water encountered in the mines may come from several different sources. In mines with light cover and without a firm solid roof, rain may seep directly into the workings. This is also true of deeper mines where pillar falls have broken the roof to the surface. The top under creek beds is generally faulty and broken and water comes through the roof at this point. Deeper mines are frequently overlain by water-bearing strata and pillar breaks and faulty top can admit water, while places running to the dip often accumulate water which comes up through the bottom. In all cases, except where passing under a year-round stream, the amount of water made is usually pretty closely connected with the weather-more water during the rainy season, less in the dry, although the deeper a mine is, the longer it is before changes in rainfall are felt. Another source of water, not connected with seasonal changes, is old water, accumulated over a period of time in abandoned workings of the mine, or from flooded adjacent mines. Here the water filters through the coal, the parting, or comes up through cracks in the bottom. Similarly, core holes, drilled to prospect the property, can be a source of unexpected water and modern safety requirements for dust-allaying water at the working faces can turn a pleasant working area into a soggy mess over the weekend if the system is not properly maintained and operated. Why water must be disposed of is obvious. If allowed to accumulate, sooner or later the mine will fill up, or at least it will become "roofed" in low places, blocking ventilation and traffic. It must be controlled long before it reaches this stage. Water standing in the working faces is unpleasant for the men who have to work there and also cuts down on the amount of work they can do, as they must watch where they step all the time. It is especially unpleasant in low coal.

Jan 1, 1981

By Robert B. Fulton

Staurolite, an iron aluminum silicate mineral, is used industrially as a high value-in-use sandblasting agent, as a premium-grade foundry sand, and as the source of aluminum in portland cement manufacture in areas where the aluminum constituent is not economically available from shale or argillaceous limestone. Staurolite is produced by E. I. duPont de Nemours and Company as a coproduct of titanium mineral production in Florida, where it was first recovered on a commercial scale in 1952 from the Pleistocene sand deposit known as Trail Ridge. At this deposit it comprises about one-fifth of the heavy mineral suite, the principal other heavy minerals being ilmenite, leucoxene, rutile, zircon, kyanite, sillimanite, and tourmaline. Staurolite and other nonconductive silicate heavy minerals are separated from the conductive titanium minerals on high-tension electrostatic separators. The silicates are then separated from one another magnetically, the staurolite being relatively magnetic. In the total process, 200 t of raw sand yield 7 t of heavy mineral concentrate from which about 1 t of staurolite is recovered. The staurolite concentrates have the characteristics given in Table 1. Staurolite concentrates are sold in three grades having the size distribution of subrounded grains shown in Table 2. All three grades are used in sandblasting for cleaning metal, for paint removal, and for cleaning masonry buildings. Staurolite compares favorably with lower priced abrasives, such as boiler slags and silica sand, in applications where the abrasive is used only once and is not recovered and recycled. Because it is hard, fine, clean, and dense it can reduce blasting time, and less material is required to clean a specific area. Compared to higher priced abrasives such as aluminum oxide and garnet, which are recycled, staurolite costs significantly less per kilogram and cleans at equal or faster rates. Its breakdown factor, which is a measure of recycling performance, is intermediate between those for aluminum oxide and garnet. [ ] [Tables 3 and 4] show the blast cleaning performance of the three staurolite grades compared to other abrasives. [Table 5] compares the longevity of premium priced abrasives during recycling. [Table 6] shows a cost comparison between silica sand and staurolite. Fringe benefits claimed for staurolite's use as an abrasive are: 1. Good flow properties and freedom from the lumps or slivers found in silica sand or boiler slag abrasives. 2. Low dusting affords good visibility, permits faster working, and reduces cleanup problems. 3. Heated storage is not required during freezing conditions, because the mineral is nonhygroscopic.

Jan 1, 1994

By Randolph E. Scheffel, David C. Fitch, David K. Hogan, Mark R. Welsh

In Oct. 1981, Ranchers Exploration and Development Corp. began producing silver bullion from the Escalante mine and mill complex, located in southwestern Utah. Current in-place reserves are estimated at 2.42 Mt (2.67 million st) of ore, containing about 936 t (33 million oz) of silver at an average grade of 428.5 g/t (12.5 oz/st). The vein-type deposit is being mined underground through a decline that is currently 864 m (2,835 ft) in length, but will ultimately be at least 1.5 km (5,100 ft) long. The mine is completely mechanized with rubber-tired equipment, and is producing 680 t/d (750 stpd) of ore, five days a week, or 177 kt/a (195,000 stpy). Additional capacity is available. The cyanide leach mill and associated refinery are presently producing 4.35 t/month (140,000 oz/month) of silver bullion-an an¬nual rate of about 52.25 t (1.68 million oz). With minor modifications, the mill can produce up to about 70 t/a (2.25 million oz/year), and expansion is currently underway. History The Escalante property is located 68 km (42 miles) west of Cedar City, UT, on the west side of the Escalante Valley, where the vein forms a gentle ridge in an area of low, rolling hills. For years, the vein has been the object of considerable exploration, but never became more than a sporadic producer, primarily due to low silver prices. Heber Holt of Enterprise, UT, staked the original four claims in 1896; they were patented in 1910. During the next 48 years, limited exploration and mining activity was conducted on the property. Systematic evaluation and development did not, however, begin until March 1958, when Samuel S. Arentz, a prominent mining engineer and consultant, formed several joint ventures with mining and oil companies that conducted drilling, drifting, mining, and water drawdown tests on the property. However, it remained uneconomic, principally because of low silver prices and costs for pumping large flows of underground water. In 1975, Ranchers obtained an option on the property and in 1979 began a drilling program. As of May 1982, the company had drilled 70 core holes, bringing the number of exploratory holes drilled on the property to 138. Test mining began in 1979 as silver prices increased, and full mine development began in 1980. Mill construction began in Sept. 1980, and on Aug 29, 1981, the first ore was fed to the mill.

Jan 9, 1982

By Alan T. Nelson, Ronald D. Mariani

The Tilden concentrator was originally constructed in 1974 for the processing of low grade hematite. Flotation is a key concentrating process in the operation, and therefore flotation optimization has been a major concern since the plant startup. This paper will discuss some of the modifications made in the process to achieve optimization. The conversion that now enables the plant to process magnetite as well as hematite also will be discussed. INTRODUCTION The Tilden Mine, located in Ishpeming, Michigan, is managed by Cleveland-Cliffs Iron Company and is owned by Cleveland-Cliffs and two Canadian steel makers. It was constructed and started production in 1974 and expanded in 1979 for processing a low grade, fine-grained hematite iron formation for which there was previously no technology available. Several papers have been written describing the unique character of the concentrating operation. The hematite plant consists of 12 grinding and concentrating lines, which utilize two stages of autogenous grinding with crushing of an intermediate sized product. The ground product is deslimed in a highly dispersed alkaline environment to reject silica. Corn starch, the flocculant, has a high affinity for iron at the operating pH range of 10.5 to 11.0. The ore is upgraded from 35% to 45% iron in the desliming step. The deslime thickener underflow is the feed to flotation. Each flotation line consists of a bank of 10 rougher cells, followed by four scavenging banks incorporating 5, 4, 3, and 3 cells each. The Wemco flotation cells are each 14.1 cubic meters (500 cubic feet) in volume. A total of 12 flotation lines are available. [ ] A reverse flotation process is employed, using a straight-chain, 8 to 10 carbon primary ether amine silica collector. At the levels used in the circuit, the amine also serves as a frother. A final concentrate is produced in the rougher cells. The concentrates from the four scavenging steps are combined and recirculated to the feed end of the roughers, with the fourth scavenger froth being discarded as a final tailing. Upgrading to about 65% iron is achieved, with control of the circuit based on final concentrate silica grade. Starch is again used, this time as a depressant for iron. During the years of operation, several key areas of the flotation process have been targeted for optimization. The most significant improvements have been realized in the following areas: *Grade Control Automation *Control of Water System Chemistry *Flotation Spray Water This paper will focus on these areas, but other minor improvements also will be briefly discussed.

Jan 1, 1993

By A. Wahab Khair, Negendra P. Reddy

INTRODUCTION The continuous mining machines which were introduced in the 1950's now account for more than half the production of coal from underground mines. Unfortunately, these continuous miners which were designed for increased productivity have also increased the concentration of respirable dust in mines. Inhalation of this respirable dust by miners causes the Coal Worker's Pneumoconiosis (CWP). Concern for CWP increased in the United States following a 1963-65 survey that showed definite x-ray evidence of CWP in 9.8 percent of working miners and in 18.2 percent of nonworking miners in the Appalachian bituminous coal region (Breslin and Niewiadomski, 1982). One of the major objectives of the Federal Coal Mine Health and Safety Act of 1969 was the prevention of CUP. This law initially limited the daily average exposure of coal miners to 3 mg/m3 of respirable dust. However, following the results of a probabilistic model developed by Jacobsen, et. al. in 1971, the limit was reduced to a mass concentration of 2 mg/m3 unless quartz content exceeds 5 percent, a rare occurrence in mines. Based on x-ray categorization prior to death it is believed that the mass concentration in the respirable fraction is a parameter of primary interest for hygenic evaluation of dust exposure. Elorgan, Burgess and Jacobson (1973) suggested that the quantity of inhaled dust, although influential is not the only aspect of exposure which is significant in producing CWP. Miners working in higher rank coal (anthracite) mines are more susceptible to CWP (Morgan, Burgess and Jacobson, 1973). This also confirms the findings of the British (Bennett, Dick and Kaplan, 1979 and Hurley, et. al., 1979) and Germany (Reisner and Rubock, 1977). Based on the cy to toxicological evidence Reisner and Rubock (1979) suggested that when coal properties are similar, increasing quartz and mineral content increases toxicity. David, Ottery and Le ?oux (1975) examined that as the severity of the CWP increased the retention of the coal dust content in the pulmonary region remained the same, while that of quartz content increased significantly . Leiteritz, Baur and Bruckman (1971) studied the various components of coal dust and showed that coal particles comprised a smaller fraction of the mass of fine dust than did particles , of quartz and minerals. It is ironic that quartz and mineral matter both comprise smaller fractions of the higher rank coal dust (Dodgson, et. al., 1971 and Dwiggins, 1981). Results of a recent research revealed that quartz or any other mineral examined had no strong or consistent effect on the incidence or progress of CWP (Attfield,

Jan 1, 1986

By John A. Haggman

Philippines is a country that is richly endowed in mineral wealth. In 1988, Climax Arimco Mining Corp. (CAMC) geologists found the Dinkidi porphyry gold-copper deposit in Nueva Vizcaya province. It is the most significant mineral discovery in that country in recent times. To date, the company has defined indicated resources of 124 Mt (136 million st) of 1-g/t (0.029-oz/st) gold and 0.4% copper. This includes a high-grade core of 7.9 Mt (8.7 million st) of 6-g/t (0.175-oz/st) gold equivalent. Total contained metal is 124.4 t (4 million oz) of gold and 499 kt (1.1 billion lb) of copper. Dinkidi's discovery resulted from initial geologic reconnaissance and recognition of the intrusive complex that is exposed at surface and anomalous in its topographic expression. This was followed by systematic exploration. This includes extensive gradient array induced polarization (IP), ground magnetics, geochemical sampling and diamond drilling. This later defined a further 16 satellite porphyry targets in the Didipio region. Climax was been granted a Financial and Technical Assistance Agreement contract with the Philippine government. It allows 100% foreign ownership of mining projects. After four years of negotiations, Climax was the first company to have such an application granted. Some 120 applications from numerous international mining companies are waiting for approval following implementation of new Philippine Mining Act of 1995. Exploration history Didipio is located 200 km (125 miles) north of Manila at an elevation of about 700 m (2,300 ft) in the remote Sierra Madre mountains. This region remains mostly unexplored. The property is accessed by 25 km (15 miles) of logging track from the town of Cabarroguis, Quirino. The area experiences an average rainfall of about 3,000 mm (118 in.) a year. Most of it falls during the July-to-November typhoon season. Didipio was first recognized as a gold province in the 1970s when native Ifugao and Ilongot hunters discovered alluvial gold in the region. The region had been prospected over the years by local companies. They were mostly prospecting for alluvial gold. However, in 1988, Climax Mining geologists identified the potential of an obscure razorback ridge protruding from the alluvial valley floor. This ridge, later named Dinkidi by Climax, was the site of several small- scale tunnels. Locals had mined narrow quartz stockwork veining and sluiced soft argillic alteration zones, using irrigation canals and sluice boxes to catch free gold.

Jan 1, 1997

By H. Maleki

Interwest Mining Company has been collecting geologic data within the last decade in preparation for underground mining of extensive coal reserves in its Bridger Coal property in southwestern Wyoming. During a preliminary data collection program in 2000, Interwest Mining completed geologic investigations, including surface mapping, drilling, aerial photography, rock testing, and limited groundwater investigations. Core from 12 exploration boreholes in the area of interest was tested to determine the mechanical properties of the coal seam and the immediate roof and floor for evaluation of the feasibility of underground mining. Geologic investigations identified the presence of one 7-to-17-ft-thick coal seam throughout the reserve under 1400 ft of cover. During a geotechnical feasibility study in 2002, the authors reviewed available data while completing comparative analyses identifying critical deficiencies in these data. The study projected moderate-to-difficult mining conditions, considering the weak nature of near-seam strata and depth of cover, and thus recommended limiting the longwall reserves to an overburden depth of less than 900 ft. These shallow remaining reserves were located at depths within regional mining experience to justify longwall mining methods. Based on a critical examination of available data, Bridger proceeded with more detailed data collection to improve understanding of the depositional and structural setting of the reserve, hydrogeologic conditions, overburden strength, and premining stress conditions. Additional data was collected during 2003 and 2004 field programs. Since 2003, Bridger has had the opportunity to observe ground behavior in a series of up to 1200-1600-ft- long rectangular openings created during highwall mining. These narrow unsupported openings remained stable in the short term in large areas but the overburden subsided locally in the long-term depending on geologic and panel layout designs. These localized subsidence events allowed an examination of in-situ coal strength based on back analyses of stress. The unextracted highwall areas near Ramp 17 provided an ideal location to access the underground reserves using a 6-entry set of mains. The existing north-south oriented longwall layout for the Bridger Coal was selected by Bridger to maximize resource recovery from the D41 reserves while considering operational and geotechnical constraints. At 45-degrees deviation from preexisting joints and mild horizontal stress field, this orientation was considered beneficial for overall stability while using both sets of joints for facilitating the caving process. The present layout utilizes pillars on 80- to 90- by 200-ft centers, 17-ft entry spans and 750-ft wide longwall panels (Figure 1). Gate pillars are on wide spacings and are intended to act as abutment-abutment pillars to limit strata deformation over the gateroads during longwall mining.

Jan 1, 2008

By James M. Barker, George S. Austin

Stone, one of the oldest building materials, today remains a well-established material throughout the construction industry. The use of natural stone is much less prevalent now than in the past. It is still widely considered to be the most aesthetically pleasing, prestigious, and durable building material. New and re-opened quarries are coming onstream to meet increased demand related to new building technology and increased residential use of stone. CLASSIFICATION No classification can completely eliminate overlap between dimension stone, aggregate, and decorative stone because most stone is multi-purpose. Many used for decorative purposes are not produced specifically for that end use. Rock otherwise considered waste in dimension stone or aggregate quarries can be decorative stone coproducts (Fig. 1). Many uses require a compromise between decorative and structural qualities (Bowles, 1992, written commu¬nication). Shipley (1945) used decorative stone interchangeably with or¬namental stone. Gary et al. (1972) defined decorative stone as that used for architectural decoration, such as mantels, columns, and store fronts, but added that it is sometimes set with silver or gold in jewelry as curio stones. Bates and Jackson (1987) also restricted decorative stone to that used for architectural decoration. Meanings of otherwise identical terms used in the stone industry differ be¬tween geologists, engineers, and quarriers. They often carry a much broader meaning for quarriers and engineers compared to their very specific use by geologists (Makens et al., 1972). Decorative stone, including ornamental stone, is more broadly defined by geologists as any stone used primarily for its color, texture, and general appearance. It is not used primarily for its strength or durability, such as construction stone, or in specific sizes, such as dimension stone. The decorative stone industry uses a much wider range of stone types compared to stone that is dimensioned. Decorative stone usually serves some structural pur¬pose, but is not load-bearing to any great extent. Weak or costly stones serve in decorative, not structural, applications. STATISTICS AND END USES Decorative and dimension stone data are difficult to separate because the US Bureau of Mines keeps statistics only on dimension stone and crushed stone. The value of domestic dimension stone production in 1990, which includes some decorative stone, was about $210 million compared to imports of about $524 million and exports of about $35 million. Production was 1 080 t of which at least one-third was for decorative uses (Taylor, 1992). The principal uses are rough blocks in building construction (23%) and monu¬ments (18%); the remainder is used as ashlar (18%), curbing (12%), and miscellaneous (29%). Major rock types are granite (50%), limestone (30%), sandstone (10%), slate (3%), marble (2%), and other (5%) (Harben, 1990). Crushed stone valued at $5.6 billion was produced in the United States in 1990 by 1700 companies operating 3400 active quarries in 48 states (Tepordei, 1991). About 52% is used in con¬struction, 9% in cement and lime manufacturing, 2% in agricul¬ture, 2% in industrial uses, and 35% for unspecified uses including decorative aggregate. Limestone and dolomite comprise about 71%, granite 14%, and traprock 8% of the stone crushed in the United States. The remaining 7% are, in descending quantity, sandstone, quartzite, miscellaneous rock, marble, shell, calcareous marl, volcanic cinder and scoria, and slate. The basic types of decorative stone are: rough stone, aggregate, cut or dressed stone, and manmade stone [(Table 1)]. Rough Stone Rough stone is used as it is found in nature with very limited processing such as minor hand shaping, edge fitting, and size or quality sorting (Perath, 1992, written communication). This stone type is often marketed locally in relatively small tonnages and includes fieldstone and flagstone. The primary end uses of rough stone are landscaping, edging, paving, or large individual stone landscape or interior accents [(Fig. 2)]. Fieldstone: Fieldstone is picked up or pried out of the ground (gleaned) without extensive quarrying and includes garden or large landscaping boulders (Austin et al., 1990, Hansen, 1969). Boulders and cobbles may be split or roughly trimmed for use in rubble walls and veneers, both interior and exterior. Popular fieldstone rock types include sandstone, basalt, limestone, gneiss, schist, quartzite, and granite, but many others are suitable. Much fieldstone is col¬lected by individuals or small companies because the industry is labor intensive and markets are small. The stone may be sold locally in small quantities from the back of vehicles (Austin et al., 1990). Fieldstone includes many rock types, sizes, and shapes with the only common denominator that it must be set by hand and be durable (Power, 1992, written communication). Moss Rock. Moss rock is fieldstone partially covered by algae, mosses, lichens, and fungi that give the rock an aged and variegated patina (Austin et al., 1990). The plants are supported by moisture and nutrients in the stone. Moss rock is used for landscaping, walls, and fireplaces. Although almost any durable rock can be a moss rock, most are slabby or rounded sandstone and limestone (Fig. 3). Flagstone: Flagstone or flagging consist of thin irregular slabs used for paving, walkways, and wall veneers. Random-shaped flagging is produced widely in the United States. Suitable stone

Jan 1, 1994

By M. O?Driscoll

The tide has turned. Developments in China that have continued to unfold over the last four years or so have now reached a point where the supply market dynamics of certain industrial minerals have changed significantly. No more is this sea change better exemplified than by the supply market for magnesite, and in particular, its processed derivative grades of dead burned magnesia (DBM), caustic calcined magnesia (CCM), and fused magnesia (FM). But what has compounded the barrage of influencing factors impacting this sector, which include the widespread effects of energy and freight cost increases, has been rising demand from certain end use markets. Clearly, the magnesia market is experiencing bittersweet conditions. But those western producers which have monitored and responded to this market transformation, stand to taste success in the near and medium future. REVOLUTION IN WORLD SUPPLY STATUS In essence, while China, as host to a wealth of magnesite resources in Liaoning province, remains a huge producer of magnesia, its supply dominance on the global magnesia market has been considerably weakened. In future, it will mainly aim to supply domestic, and to some extent east Asian markets. Magnesia consumers outside China are now scrabbling for alternative, ?western? sources of supply. At the same time, magnesia producers outside China that weathered the storm of low cost Chinese DBM and FM imports flooding their regional markets for the last 20 years (and many did not), are busy investing in production capacity increases to meet not just demand from falling Chinese supply, but also demand from end use markets which is picking up. And it is not just a matter of expanding capacity of existing magnesia product lines. In the face of plentiful lower cost DBM and FM substitutes from China over the last couple of decades, many western magnesia producers started to switch their focus to the non-refractory markets (eg. environmental, specialities) which promised lucrative opportunities (and still do) and crucially, a lessening of sales revenue dependence on the DBM refractories market. Now we have the likes of Baymag (Canada), Bommag (Serbia/Turkey), and Kumas (Turkey) not only expanding capacities but seriously considering diversifying into DBM and FM production. Perhaps DSP (Israel), having ceased its famous high purity DBM activity as recently as 2005, might even reconsider a return to the ?other side?. But it?s not all DBM and FM. The CCM and specialities markets are also demanding a response, and so Baymag, DSP, Martin Marietta (USA), Premier Chemicals (USA), Qmag (Australia), MGR (Spain), and Magnifin (Austria) are also stepping up to the market on the non-refractory side. From most accounts, the situation with regard to Chinese supply has been described as one of turmoil, with prices rising, leading to improving opportunities for western producers. The majors are conducting huge expansion programmes, such as Qmag, Magnezit (Russia), and the new look Magnesita (Brazil). Even smaller players such as Causmag (Australia) and Dalmia (India) are looking to increase production. In short, the western magnesia supply sector is witnessing a landmark event in capacity investment and market share penetration as China?s previously dominant role begins to weaken (for a review of western capacity expansions see Industrial Minerals, September 2008, p.28). MERGER & ACQUISITION ACTIVITY Another response of course has been a raft of mergers and acquisitions in order to secure resources and supply outside China. These have included private equity groups taking over Magnesita and Qmag; Imerys entering the FM business through UCM; Peñoles further consolidating the electrical grade FM market by acquiring Minco; Martin Marietta acquiring Morton Salt?s Specialty Magnesia Group; Bomex entering the market through Calmag (now Bommag); LWB acquiring Magnesita; and Magnezit?s pursuit of Slovakian magnesia, with Slovmag aboard while currently wooing SMZ. Interestingly, RHI has gone ?the other way?, and has invested in a state of the art joint venture operation at Dashiqiao, Liaoning, with which it aims to secure high quality feedstock for its Chinese refractory plants. Tata Refractories Ltd has announced intentions to follow suit, and other majors, such as Qmag are known to be interested in seeking options. Naturally, there is now renewed interest from several parties in the idled 50,000 tpa DBM/10,000 CCM Jormag facility, Jordan, and the commercially undeveloped Zhargat project, Saudi Arabia. Elsewhere, there will no doubt be suitors for stakes in SMZ, in Slovakia, and Magnohrom, in Serbia, whose respective ownerships are on the block. THE ?CHINA FACTOR? The ?China factor? has been key to magnesia?s market change, and has significantly influenced prices and availability of material to global markets. Resources & production China hosts the largest share of the world?s magnesite deposits, accounting for 26% or 3,319m. tonnes of predominantly sparry magnesite (followed by North Korea, 24%, and Russia, 22%; Wilson 2008). China accounts for 44% of total world magnesite mined (19m. tonnes in 2007) and 50% of the total world magnesia produced (8.2m. tonnes CCM, DBM, FM; Wilson 2008). Russia follows China at 13% of magnesia supply, illustrating China?s dominance of magnesia production. Apart from one producer in Shandong province, central eastern China, the country?s magnesite resources and production are concentrated around the cities of Haicheng and Dashiqiao in Liaoning province, north-east China. Although there are also producers in the Xiuyan district of south-east Liaoning, the Haicheng-Dashiqiao Magnesite Belt hosts the majority of producers. In total these may number 200-300 of large,

Jan 1, 2009

By Susanne Pignolet-Brandom

Flash smelting products of a copper sulfide concentrate collected at successive levels in an experimental reactor have been studied by reflected light microscopy, scanning electron microscope - energy dispersive analysis and electron nicroprobe analysis. Chalcopyrite, intermediate solid solution, bornite, chalcocite, metallic copper, magnetite, iron-copper spinel, pyrrhotite, hematite, delafossite and cuprite have been identified in the product samples. The detailed reflected light microscopy study of these phases and their abundance in the samples has resulted in the identification of the series of reactions during flash smelting.

Jan 1, 1985