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By G. V. Sullivan

The Federal Bureau of Mines examined tailings samples from 13 western copper operations to determine the potential for recovering rutile. Results showed that six of the samples contained sufficient liberated titania to be considered potential sources. One sample contained 0.75 percent Ti02 or 0.40 percent liberated rutile. Batch flotation studies developed a procedure that recovered 70 percent of the liberated rutile at a grade of 34.7 percent TiO2. Sulfuric acid leaching studies on 12 of the samples dissolved 100 to 2,200 grams Cu per ton of tailings; leaching of the slimes showed slightly more copper was dissolved, 1,800 to 3,300 grams Cu per ton of minus lo-micrometer material. Bulk sulfide flotation of nine of the samples resulted in concentrates that contained 0.38 to 2.0 percent Cu and 0.7 to 2.1 and 6.9 to 42.9 grams per ton Au and Ag, respectively.

Jan 1, 1981

By J. A. Botín

In modern mining, the core of mine planning and design is the block model, a simplified representation wherein the ore deposit is divided in a number of virtual rock cubes, each characterized by technical and economics parameters. Mining engineers use this model to design the mine and develop production schedules. Traditionally, sustainability issues are not considered in the mine design, planning and engineering processes but rather are incorporated in the later stages of the feasibility evaluation as necessary expenditures and operating costs. This paper describes the results of a research project aiming to develop a methodology to incorporate sustainability parameters directly into the block model, so that they become an integral part of the design, planning and scheduling processes. It is considered that the incorporation of sustainable parameters into the block model would yield a more responsible mine design and improve mining projects sustainability performance

Feb 27, 2013

Effective tools are not only essential for the design of practical and efficient underground mine ventilation systems but must also play an integral part in the operation of the mine. One such tool is VUMA-coal ventilation simulation software. VUMA-coal is Windows-based and has been specially designed for the simulation of underground coal mines and includes user-friendly coal specific interfaces and three-dimensional graphics designed to facilitate in the construction and analysis of networks. VUMA-coal can be utilised during the design phase of new collieries, be used to conduct ?what-if? and optimisation studies, and be integrated into the ?day-to-day? ventilation related operations. In addition, VUMA-coal can be used to simulate dust and gas distribution in a ventilation network, as well as predict the impact of control measures. This paper describes the interactive VUMA-coal ventilation simulation software and its underlying design features and outlines a case study of how VUMA-coal has been integrated into Anglo Coal?s Ventilation and Occupational Hygiene Engineering [VOHE] department as part of the standard planning and evaluation process in their operational mines as well as long term planning and future mining operations.

Jan 1, 2009

By Frank P. Frandina, Michael J. Loehr, Paul E. Gabryszak

The Heartland Corridor Clearance Improvement Project is part of a regional Public-Private initiative to allow double stack container trains a direct route from the Port of Norfolk through Virginia, West Virginia, Eastern Kentucky and into Central Ohio. In the Preliminary Engineering stage, 30 existing tunnels were assessed and 28 were found to need clearance improvements to allow the use of double stack trains. Alternative tunnel enlargement methods were developed to achieve the desired clearance improvements, while maintaining structural integrity and minimizing disruption to train traffic. This paper summarizes the geometric and condition investigations of the existing tunnels, as well as describes the preferred methodologies to achieve the desired clearance. The clearance improvements will be implemented beginning in the summer of 2007.

Jan 1, 2007

By Dewa Putu Agus Sudarmaja, Widi Astuti, Januarti Jaya Eka Putri, Yayat Iman Supriyatna, Himawan Tri Bayu Murti Petrus, Siti Nurul Aisyiyah Jenie, Ferian Anggara, Anisa Maulidia, Vincent Sutresno Hadi Sujoto, Sutijan, Wayan Warmada

With the rapid development of renewable energy technology, such as electricity reserves to store energy for electric and battery-electric vehicles, environmentally friendly energy sources are expected to continue to be tapped to meet mobility and stationary energy needs. Lithium is the most attractive candidate for energy storage development. The presence of lithium in the Sidoarjo mud in Indonesia makes it interesting for use as an alternative source of raw material. In this research work, sulfuric acid was used as a solvent, and hydrogen peroxide as an oxidation agent. Lithium recovery was studied with a leaching process at various temperatures (30, 45 and 60 °C) and sulfuric acid concentrations (0.5, 1, 2, 3, 4 and 6 mol/L). Increases in temperature, sulfuric acid concentration and hydrogen peroxide concentration were found to directly affect the lithium recovery. An increase in sulfuric acid concentration is directly proportional to the recovery of lithium in the solution. Optimum conditions were obtained at a sulfuric acid concentration of 6 mol/L and temperature of 60 °C. The shrinking core model was used to study leaching kinetics, and ash layer diffusion acted as a rate-controlling process. The activation energy of the leaching process was found to be 38.90 kJ/mol.

Jun 1, 2024

By M. L. Chapman

When originally accepting the assignment, the Client had described the property as vacant land. A prior appraisal was later discovered and revealed that the property included a mineral rights appraisal. Further research disclosed an old dispute with a court ruling based on testimony by AIMA-certified experts. The present appraisal was therefore based on data from findings of the court with a fresh look at discount rate, absorption period, and residual land value. Fortuitously, the property subsequently sold on the market, while meeting all criteria for Fair Market Value. This provided insight into a market-based discount rate. Also, the absorption period could be adjusted in accordance with the current market conditions. The residual land value was minimal (based on comparable sales and projecting trends in market) as the property was affected by a flood zone and the surrounding development was sparse. PROPERTY HISTORY: A GREAT INVESTMENT Briefly, the subject property is located in the floodplain region of the San Jacinto River, a short river in southeastern Texas that flows into Galveston Bay, filled with sand and gravel. This location is on the north side of Houston, north of Beltway 8 on the east side of I-59.

Jan 1, 2012

By B. Seegmiller, W. A. Hustrulid, O. Stephansson

Introduction The concept of in-the-wall haulage was formally introduced to the mining community in the early 1970s. The basic concept is simple. Rather than the traditional in-the-pit haulage/access ramp, single or twin ramps driven in the pit wall are used with crosscut connections on the working levels. The in-the-wall pit haulage system could be considered as part of the opening of a new mine, a way of completing to final depth an existing pit, or part of the transition from open-pit to underground mining with the ramp remaining as part of the access. This paper is intended to acquaint the reader, particularly mine planning engineers, with some of the possibilities presented by the concept. Nordstrom (1983) said, "In many countries, mining engineers are either open-pit or underground specialists and, hence, often don't consider the possibilities for combining the best aspects of each into a hybrid mining system." This is believed to be true. There are few "hybrid mines" employing, for example, open-pit methods with underground haulage systems. One reason is the variance with the experience of the mining company and the engineers involved. However, with today's low metal prices, innovative and unconventional thinking may be just what a property requires to remain in production. All the possible situations where an in-the-wall ramp might be considered cannot be presented here. Rather, a few examples will be discussed for illustration purposes. Potential savings in stripping costs A major cost in open-pit mining is associated with the removal of overburden and waste rock. Although much of this stripping cannot be avoided, the volume (V) associated with the presence of an in-pit haul road can be quite large, V 100 (pit depth) a (road width) 2 x road grade (%) As an example, consider a vertical ore body 400 m (1312 ft) long, 60 m (197 ft) thick, extending to a depth of 160 m (525 ft). Cross sections are shown in Fig. 1. Other data are: ore tonnage = 11 Mt (12 million st), ore density = 2.9 t/m3 (0.09 st per cu ft), waste rock density = 2.7 t/m3 (0.08 st per cu ft), road width = 20 m (66 ft), road gradient = 10%, and level interval = 10 m (33 ft). The evaluation has been made assuming that the pit slope (0) is constant around the pit except on the side with the haul road. On this side, the slope is decreased to allow for the road. Slope angles of 30° to 60° have been selected to indicate the dependence of stripping volume (V) on slope angle. The stripping at the ends have been calculated assuming representation by cones. The volume associated with the haul road is 2.6 Mm3 (92.8 million cu ft). A comparison of the strip volumes involved is shown in Table 1. As can be seen, the percent savings in strip volume varies from about 4% to 17% and depends greatly on the slope angle. It is perhaps more interesting to examine the savings in stripping costs that could then be used to cover the cost of the access ramp and the crosscuts for stripping costs of $3/m3 ($0.08 per cu ft). The cost with the road volume is 2.6 million x 3 = $7.8 million.

Jan 1, 1988

By Cameron Dunn, Callum Fryer, Carmen Hu, Craig Sewell, Peter Stakne, Andrew Raine

The Woodsmith Project is located in north-eastern England within the North York Moors National Park. The strata contains the world’s largest known high-grade Polyhalite resource of around 2.66 billion tonnes. The Project was initiated by York Potash Limited (YPL), a subsidiary of Sirius Minerals plc., but in March 2020, a takeover was completed by Anglo American, one of the largest mining companies in the world. Polyhalite is an evaporite mineral and contains essential nutrients required for plant growth. The Polyhalite will be extracted via two deep mine shafts and transported underground from the mine site to a Materials Handling Facility on a Mineral Transport System (MTS), which comprises a high-capacity conveyor belt system located within a 4.9 m ID tunnel. In 2018, STRABAG were awarded with the design and construction of the MTS tunnel and associated infrastructure, including the high-capacity conveyor system. The scope of work consists of 300 m of open cut, 100 m SCL tunnel, 36.5 km TBM drive (originally anticipated as three separate TBM drives) along with a network of underground caverns at the Woodsmith Mine. The construction of two 380 m deep intermediate access shafts was later awarded as Amendments to the Contract. The tunnel is constructed using a Herrenknecht Single Shield TBM. As a result of how the various elements of the project have developed over time, the first TBM drive has been extended to ~29 km making it the worlds longest single heading TBM Tunnel when complete. The option of completing the entire tunnel length with a single heading TBM drive is currently being assessed by Anglo American and STRABAG. This unprecedented tunnel length requires two intermediate shafts to provide essential access and services, the first intermediate shaft is at Lockwood beck (LWB) and the second is at Ladycross (LDX). The LWB shaft has been constructed utilising the innovative Blind Boring technique never used before in the UK. This paper will explain the methodology and technical risk management associated with this technique. The authors believe that the collaborative approach between all contract partners is and continues to be the important contributor for the success of this project.

Jun 13, 2023

Bortech’s BoreClad corrosion protection system will be displayed at MINExpo. BoreClad is an innovative system for inside diameter cladding of nozzles, tubes and other circumferential devices. Pressure vessels, mud pumps and other mining equipment must be protected from corrosive caustic chemicals. Usually manufactured from mild steel, they can be clad with a protective material, such as stainless steel, to provide corrosion resistance. However, most welding processes leave voids where corrosion can damage the base metal. BoreClad provides void-free welds by combining gas metal arc welding with automated feed and rotation of chuckable parts. Workpieces of up to 915-mm- (36-in.-) long with a 460-mm- (18-in.-)swing can be clad with stainless steel, Iconel or other exotic metals. In addition to eliminating the voids associated with hand welds, BoreCladcan lay more than 2.2 kg/h (5 lbs/hour) of metal. That is a significant reduction in weld time. Web sitewww.bortech.com.

Jan 1, 2004

By J. J. Crowhurst

The Carson Hill mine in Calaveras County, California, was explored and developed by Grandview Resources Inc., as an open-pit heap-leached operation, with production commencing in the fall of 1986. Grandview entered the district in 1983 after a 40-year period of quiescense, the initial contact being by Barden Stevenot, V.F. Hollister, and J.J. Crowhurst. The last dividend declared by the Carson Hill Gold Mining Corp. was 3 1/2$ per share paid in January 1943. The mine was originally discovered in August 1848 by Jim Carson. Preceding the entry by Grandview in 1983, Cyprus Mines explored Carson Hill in 1975, and drill exposed a mineral inventory of 7.3 Mt (7,982,000 tons) averaging 1.4 g/t (0.040 oz per ton) gold. From that time until the Grandview entry, R.W. Thomssen and B.E. Stevenot continued investigations into the potential of the deposit. Grandview successfully investigated many of the metallurgical problems involved with heap leaching the ore. Metallurgical investigations were con- cluded under the direction of R.S. Shoemaker. By November 1984, J.A.C. Ross had prepared a prefeasibility report that included the earliest Grandview investigations. This report concluded that the deposit contained at least 6.8 Mt (7,440,000 tons) of ore that average 1.5 g/t (0.045 oz per ton) gold and that this ore could be mined with a 2: 1 stripping ratio. A final feasibility report was prepared by D.W. Pringle in September 1985. It showed a final ore reserve of 14.7 Mt (16,160,000 tons) of 1.7 g/t (0.049 oz per ton) gold. Attached is a description of the geology of the deposit from an internal report covering Cyprus activity in 1975 and a brief portion of USBM IC 6940 (1937). General location and geologic maps attached are from the Pringle report.

Jan 1, 1990

By K. Huttu, A. van Wageningen

"The Kittilä underground mine in northern Finland is extracting one of the largest known gold deposits in Europe. Kittilä achieved commercial production on May 1, 2009, becoming Agnico Eagle’s first mine to open outside of Canada. The Kittilä orebodies were initially mined from two openpits – Suuri and Roura – and underground operations were added in October 2010. The pits were mined out in November 2012, so mining is now entirely underground in the Suuri and Roura deposits. Additional, small openpits will mine any remaining near-surface reserves in the future.The Kittilä Mine is located in the Lapland region of northern Finland, approximately 900 km (560 miles) north of Helsinki and 150 km (93 miles) north of the Arctic Circle. With a mine life estimated through 2034, Kittilä is Agnico’s longest-life mine.The mine’s proven and probable reserves contain 32 Mt (4.7 million oz) at 4.6 g/t (0.13 oz/ st) gold. Ore has been mined from underground since 2010. The original 3 kt/d (3,300 stpd) operation underwent an expansion project to increase the throughput capacity of the mill by 1 kt/d (1,100 stpd), starting by the end of 2014.The region around the Kittilä Mine is underlain by mafic volcanic and sedimentary rocks of the Kittilä Greenstone Belt oriented nearly vertical. This greenstone belt is similar to those hosting Agnico Eagle’s Canadian deposits in Quebec’s Abitibi region and Nunavut. At the contact point between ironrich and magnesium-rich rocks lies a 100- to 200- m (330- to 660-ft) thick structural zone known as the Suurikuusikko Trend. This trend hosts the Kittilä deposit, where multiple mineralized zones have been traced over a strike length of more than 25 km (15 miles).Most of Agnico Eagle’s work has focused on a 4.5-km (2-mile) segment of the Suurikuusikko Trend that hosts the five main zones of the known gold reserves and resources. From south to north they are: Ketola, Etela, Suuri, Roura and Rimpi.The ore at the Kittilä Mine is refractory, making gold extraction relatively difficult because the gold is generally locked inside the two main sulfide minerals – arsenopyrite and arsenic-rich pyrite. Only about 4 percent is very fine-grained “free gold” (not locked in sulfide grains) and easier to extract. Most of this free gold is found in the outer, oxidized or eroded sections of the ore."

Jan 1, 2015

By Pelin Altinkaya, Maria Leikola, Jussi Liipo, Mika Haapalainen, Mari Lundström, Eero Kolehmainen

With decreasing ore grades, the tailings of mining operations are becoming of increasing interest as metal-containing secondary raw materials. The objective of the current work was to investigate chloride leaching of gold, copper, cobalt, nickel, and zinc present in the flotation tailings. In the current study, the effect of cupric ion as an oxidant (0–50 g/L) and NaCl (150–250 g/L) on metals extraction was investigated. The other parameters, such as pH (1.8), temperature (95 °C), solid/liquid ratio (25%), oxygen feed rate (1200 mL/min), leaching time (24 h), and stirring rate (950 rpm), were kept constant. Gold dissolution rate was shown to increase with increase in cupric ion concentrations up to 50 g/L. Also, increase in NaCl concentration up to 250 g/L increased gold extraction. Majority of the copper present in the flotation tailings could be dissolved with all the all solutions investigated. The other base metals, cobalt, nickel, and zinc extractions were also shown to increase with increase in the cupric ion and sodium chloride concentration. However, even in the absence of cupric ion addition (t = 72 h, 250 g/L NaCl), the final extraction of Cu, Ni, Zn, Co, and Fe increased up to 98, 93, 83, 76, and 80%, respectively. This shows the power of inherently originating oxidants present in the tailings. Furthermore, the solid analysis of the leach residue indicated that leaching the flotation tailings in pure NaCl may result in partial gold dissolution. The results demonstrate that significant amount of metals present in the tailings could be extracted even with only NaCl as added chemical in the presence of oxygen feed, oxidizing agents originating directly from the raw material. This can provide an advantageous cyanide-free method for extraction of metals from very low-grade tailings with low chemical consumption.

The Río Agrio of northern Patagonia, Argentina, is highly acidic due to inputs of volcanic gases (SO2, HCl, HF) in its headwaters. The watershed includes a hyper-acidic (pH <0.5) volcanic crater lake, a larger glacially carved lake that is also strongly acidic (pH 2.6 to 2.7) and an acidic river that flows for more than 40 km (25 miles) before its pH is completely neutralized. This unique watershed has many geochemical and biological similarities to acid waters caused by mining. This paper describes the Río Agrio watershed and draws a comparison with some famous acid mine waters. Learning more about naturally acidic lakes and river systems could have important applications to the bioleaching of ores or biological treatment of mine waters. Therefore, it should be of interest to the general community of mining and metallurgical engineers.

Jan 1, 2008

By Syd S. Peng

Introduction The most commonly used ground control design in coal mining is the determination of pillar size by various pillar strength formulae for mine layout (Bieniawski, 1983; Holland, 1973; Hustrulid, 1976; Obert and Duvall 1967; Wilson and Ashwin, 1972). However, there have been few documented case histories concerning the validity of those strength formulae. Recently there are indications that those strength formulae tend to overdesign considerably. From a safety point of view, there is nothing wrong with overdesign. But an overdesign with large chain pillars reduces the flexibility of machine movement and decreases production. It also reduces coal recovery and, thus, shortens the reserve&apos;s life. This is especially true for small operators who own a limited reserve. Those small mines, through past practices or out of necessity, experiment with smaller pillars than those predicted by the strength formulae. Conclusive ev¬idence has been collected that points to the need of a new concept of coal pillar design. Two case studies will be cited to illustrate this point. 3-D US pillar strength formulae Figure 1 shows a sectional layout of a one-unit mine. The seam was about 1.3 m (50 in.) thick with a uniform cover of 64 m (210 ft). A house was located near the center of the section. The homeowner demanded that no surface subsidence and domestic well dewatering should ever be allowed as a result of underground room-and-pillar mining. Physical properties - strength and Young&apos;s modulus - of the rock strata including coal were obtained by testing small samples in the laboratory. The uniaxial compressive strength of the coal was 3660 kPa (823 psi). Using various pillar strength formulae, the homeowner recommended that square chain pillars be at least 11 m (35 ft) wide. The coal company, however, desired to pursue optimum coal recovery in the safest and most economical manner. They engaged in three-dimensional finite element analysis of the section taking into account the complete stratigraphic sequence of the overburden. Results indicated that the section can be developed by 5.5 m (18 ft) wide entries at 21 m (70 ft) center-to-center. In retreat, the pillars can be splitted along both centerlines leaving a stump pillar of 5.5 m (18 ft) square at each of the four corners. The 5.5 m (18 ft) square stump pillars still maintained a safety factor of 1.5, although the safety factors calculated by various pillar strength

Jan 11, 1986

The Mine Safety and Health Administration (MSHA) conducted a respirable dust survey to evaluate the effectiveness of a water curtain utilized for dust control on continuous miner sections. The purpose of this survey was to determine if the water curtain was effective in reducing respirable dust concentrations, since several mines currently use water curtains to control miners? respirable dust exposures. The survey was conducted on a supersection of an underground coal mine, which utilized two continuous miners on the same split of air. Each continuous miner alternated taking a cut so that both continuous miners would not operate at the same time. The underground coal mine is located in Knox County, Indiana. The water curtain evaluated in this survey was located in the last open crosscut between the intake side continuous miner and the return side continuous miner. When the intake side continuous miner cuts coal, its exhaust air would be used to ventilate the return side of the section. The intent of the water curtain was to reduce the intake air respirable dust concentration (generated by the intake side continuous miner) before the air stream ventilated the return side of the section. To determine the effectiveness of the water curtain in reducing respirable dust concentrations, respirable dust sampling pumps were placed on both the intake side and the return side of the water curtain. Additional respirable dust sampling pumps were placed farther downstream from the water curtain to monitor dust levels downstream of the water curtain, but in an area where respirable dust would not be created by anything other than the intake side continuous miner or additional section air. The water curtain and the respirable dust sampling pumps were operated only when the intake side continuous miner was cutting coal. The samples were analyzed for respirable dust and quartz content. The water curtain showed minimal effect in controlling respirable dust and respirable quartz dust, but the reduction was so negligible that this survey could not quantify the reduction. The measured reductions are within the accuracy of the sampling equipment.

Jan 1, 2009

By William Reininger

Microtunneling provides a necessary solution for multiple utility crossings associated with the New York Metropolitan Transportation Authority Capital Construction Company?s (MTACC) Eastside Access Project for the Long Island Rail Road. Construction of four TBM tunnels and their approach structures at the eastern end of the project in Sunnyside, Queens required significant modification and expansion of the railroad infrastructure at this location. Maintenance of railroad traffic during project construction is a principal requirement. As a consequence fourteen (14) utility crossing segments are being installed under active traffic by microtunneling methods. Microtunnels were installed in a variety of conditions including depths, types of soils and construction restraints.

By H. Ipek

Industrial grinding machines used in the mineral processing industries are mostly of the tumbling mill type. These mills exist in various types: ball, rod, pebble, autogenous, and semi-autogenous. Balls and rods are traditionally media shapes used in the tumbling mills. In resent years, as an alternative to balls, grinding charges with various shapes are being used. Cylpebs are slightly tapered cylindrical grinding charges with length equaling diameter, and all the edges being radiused. Cylpebs of equal diameter and length have 50% greater surface area and 45% greater weight than balls of the same material. In addition, they have 9% higher bulk density than steel balls, or 12% higher bulk density than cast balls. An important advantage of cylpebs over balls is that cylpebs produce narrower granulometric distribution due to their shapes. However, some discrepancies were found in the literature regarding the milling performance of cylpebs in comparison with balls (Chi, 2004). In recent studies, Chi (2004) has provided detailed information on cylpebs and has compared these to balls in relation to media mass, size distribution, surface area, and input specific energy. Ipek (2006) has investigated the effects of cylpebs and balls with equal mass and surface area on the breakage rate of quartz, using cylpebs and balls separately. Similarly, Lameck and Moys (2006) have investigated the effects of balls and worn balls on breakage rate of material using worn balls as grinding environment. However, no study has attempted to investigate the effects of grinding environment other than balls on the brekage distribution. As indicated by Prasher (1987), the breakage distribution function is independent of the following variables in ball mills: hold-up (Mika et al., 1967; Kelsall et al. 1969), mill speed and ball load (Herbst and Fuerstenau, 1972), media load and ball density (Kelsall et al., 1973), mill diameter (Malghan and Fuerstenau, 1976), and powder charge (Shoji et al. 1980). The aim of the present paper is to compare the effects of cylpebs and balls on the breakage distribution function of quartz with respect to same mass and same surface area. BACKGROUND Comminution as a rate process is called kinetic approach. Studies on the kinetic approach are reviewed by Oner (2001). Oner states that grinding kinetics are based on studies by Kolmogorov in 1941 and that Robert (1950) is the first to suggest this approach. Later, Melzak, Loeveday, and Reid have improved the grinding kinetics equation; and, Austin and Lynch have conducted two basic studies on the analytical solutions of these equations. Considering the results of studies comparing the size reduction mechanisms, it has been accepted that breakage constitutes of two operations. These two basic operations, first defined by Epstein (1948), are those of selection of material for breakage and the distribution of breakage material to particle sizes smaller than the initial size (Lynch, 1977). Epstein defined the two operations as: ? Pn(y), the probability of breakage of a particle size in the nth step of the breakage process. ? F(x,y), the cumulative distributions by weight of particles of size x<y arising from the breakage of a unit mass of a size y. Probability of breakage is called a selection or a breakage-rate function (Si). Characteristics size distribution after breakage is called a breakage distribution or appearance function (Bi,j). Due to the difficulties in representing particle size distributions mathematically and constantly, the batch equation based on particle size fraction is suggested. Thus, as is the case in breakage distribution and specific rate of breakage, the breakage parameters were defined by batch grinding experiments conducted in laboratories (Oner, 2001). When effective breakage occurs in the mill, the breakage of material with single feed size fraction usually fits first order grinding kinetics and the specific rate of breakage values can be calculated through the slope of semi-log plot. The values of the breakage distributions are deduced from the size distributions at short grinding times where there is mainly size 1 material breaking and only small amount of smaller sizes to rebreak (Austin et. al., 1984). B values can be calculated by using BII. [ ] where j is the top size of the charge, Pi(t) is the cumulative percent undersize of i th size interval at the short grinding time which gives no more than about 30% broken out of the top size interval. Cumulative primary breakage distribution, Bi,j, is also defined in an empirical form by [ ] where xi is the top size and Bi,j is the weight fraction of primary breakage products. The parameters F, ? and ß define the size distribution of the material being ground. On plotting size versus Bi,j on log scale, the slope of the lower portion of the curve gives the value of ?, the slope of the upper portion of the curve gives the value of ß, and F is the intercept as in Fig.1. [ ]

Jan 1, 2009

By E. Z. Manos

The Detroit salt mine is operated by the Detroit Salt Co. It is one of the oldest underground room-and-pillar mines in the United States still in production. Mining at the Detroit salt mine began in 1906 with the sinking of the No. 1 shaft. Mining continued through the years, becoming progressively more modern in its equipment and techniques, until it was closed in 1983. In 1997, after a period of care and maintenance, the newly formed Detroit Salt Co. decided to reopen the mine. Upon completion of an extensive rehabilitation program, salt production resumed in 1998. Production currently continues at a rate of 910kt/a (1 million stpy). To date, the Detroit salt mine has mined more than 45 Mt (50 million st).This mine is unique because it lies under a heavily urbanized area within metropolitan Detroit. The mine and its history make a good case study for the continuing coexistence between the mining industry and its environs.

Jan 1, 2003

By L. Harris

The Yanacocha project is the first foreign investment in Peruvian mining since Cuajone during 1976. Three years after the project began, it has become the largest producer of gold from heap-leach operations in the world and the top producer in Latin America. The success of Minera Yanacocha SA (MYSA) opens a new chapter in Peruvian mining with the exploitation of its widely spread gold deposits. And it has attracted the interest of foreign investors in Peru’s mineral resources. Construction began on Aug. 1, 1992. The first gold bar was poured on Aug. 7, 1993 from ore from the Carachugo Mine. During 1993, Minera Yanacocha produced 2.5 t (81,000 oz) of gold and 9.4 t (304,000 oz) during1994 (40% above budget). Production was expected to exceed 15.5 t (500,000oz) during 1995 with the start-up the Maqui Maqui Mine.

Jan 1, 1996

By Congling Shi, Marek Borowski, Yue Wang, Jianwei Cheng, Xixi Zhang, Xuan Xu

As the first barrier to prevent accidents caused by explosions in underground coal mines, the integrity and stability of mine seals have great effects on mine workers’ lives and underground production safety. In this paper, the mechanical responding performance of typical mine seals under the impact load are studied by using numerical simulation method. By such simulation studies, the safety of different seals subjected to the explosion load is evaluated. Seals’stability are also investigated under effects caused by various influence factors including thickness of mine seal, cutting depth into surrounding rocks, rock type, and the roof-tofloor convergence under given the explosion load. The research results can provide theoretical basis and reasonable improvements for mine seal’s construction practices for improving its anti-explosion performance.