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By E. C. Robertson

D.G. Wilder I found the suggestion that the amount of displacement of a fault can be numerically related to the thickness of gouge or breccia to be both intuitively satisfying and intriguing. I have long agreed that there is some type of relationship between the amount of gouge and the amount of displacement of faults. I congratulate the author for developing a numerical relationship between them. However, I am concerned that the limits for applying this relationship be fully understood. An underlying assumption in this approach is that there is either a uniform thickness of gouge or breccia along a given fault or the thickness does not vary widely. Since it is not always possible to confirm this, the displacements derived by this method should be viewed with caution unless significant fault extent can be observed. At the Nevada Test Site, in drifts constructed in granite for test emplacement of spent nuclear reactor fuel, we found a fault with 0.3 to 0.4 m (12 to 16 in.) of clay gouge. Within a few meters of this location, the fault had no clay gouge, but rather consisted of a highly fractured zone with significantly altered rock and some slickensides. Based on Fig. 1, the 0.3 to 0.4 m (12 to 16 in.) thickness of gouge would indicate a displacement in excess of 30 m (98 ft). However, no gouge thickness would indicate essentially no displacement. Based on a quartz vein that terminated on the fault, and is not identified nearby, an estimated displacement of more than a few meters was made. This estimate is consistent with that obtained using the regression line proposed in the paper if the 0.3 to 0.4 m (12 to 16 in.) thickness for the gouge is used. However, using the regression curves with zero thickness would not yield results consistent with what was observed in the field. Therefore, it is important to recognize that the suggested procedure would properly yield a range of probable displacements. ? *Work performed under the auspices of the US Department of Energy by the Lawrence Livermore National Laboratory under Contract W-7405-Eng-48. Reply by E.C. Robertson It is certainly true that the t (thickness) of gg-bx (gouge and breccia) on a fault does vary along the fault. My observations have been that near the termination of a fault, the displacement d is small and the t is also small, whereas the maximum d and t will usually be found in the central part of the fault. The information on gg-bx and t of the fault found in granite in the NTS tunnel by Mr. Wilder could be interpreted somewhat differently than he does. He speaks of the fault changing within a few meters from 0.3 to 0.4 m (12 to 16 in.) of clay gg to "a highly fractured zone with significantly altered rock and some slickensides," but no gg. The highly fractured rock may be taken to be bx, rock not so finely ground as gg but still crushed by the fault movement, equivalent to the gg in my usage, and probably occupying about the same t. Mr. Wilder's estimates for the fault in the NTS tunnel for t of 0.3 to 0.4 m (12 to 16 in.) and for d of a quartz vein, in excess of "a few meters," would place the point on the low side of the central trend line in my Fig. 1, at the lower limit. There is, of course, a problem with determining d using displacement of only one planar surface. It would be greater or lesser depending on the rake of the movement. Finally, estimating the d of a fault from its t should be made with awareness of our present uncertainties, as pointed out by Mr. Wilder. Although the central trend line in my Fig. 1 has a ratio of d/t of 100, I have put the limiting ratios at 10 and 1000. Understanding of the values of the ratio will be improved only with collection of more data, for which the discussion of Mr. Wilder is much appreciated. G.C. Waterman E.C. Robertson's paper provides significant information to a geologist attempting to deduce fault offset by noting the products of structural dislocation. However, considerable mapping in underground and open-pit mines, and examination of structures produced in different geological settings, have convinced me that gouge and breccia thickness are controlled by geological conditions and fault movement. The following paragraphs suggest geological variables that control them. 1. Depth of Loading A near-surface fault resulting from tensional stress has more breccia/gouge than is produced by a similar stress at considerable depth. A deep-loaded compressional stress may produce a linear zone of schist, or structural dislocation may occur along an earlier formed belt of schist. Such "shear zones" are common in Canadian mines in precambrian rocks. In neither case can offset be directly deduced by an analysis of the minimal gouge/breccia in the shistose rocks. At greater depth, stress may be partially to wholly relieved by flowage. I vividly recall first noting the regional "Midas Thrust" in the Lark mine, Bingham Mining District, UT (where we called the structure the North Fault). My recorded notes, as I remember them, showed a narrow gouge streak separated the `Jordan" and "Commercial" limestone units from impure, muddy limestone beds of uncertain stratigraphic position. The visible structure did not indicate the great importance of this premineral fault

Jan 12, 1984

By I. Loomis, D. Rachmad, I. Duckworth

During the past two years the Underground Mine Ventilation Group at PT Freeport Indonesia has undertaken to upgrade their Hygiene Laboratory. This process has included upgrading equipment for dust sampling and gas monitoring, as well as procurement of an OCEC analyzer for Diesel Particulate Matter analysis. While the laboratory equipment represents one factor, these changes have also required re-evaluation of the sampling protocols and the burden of ventilation monitoring personnel. This paper provides a review of these experiences.

Jan 1, 2006

By T. W. Camm, J. C. Johnson, S. D. Rosenthal

"Most engineers are bright, hard-working, reliable and prefer to avoid conflict. An engineering curriculum tends to self-select these characteristics. By most standards, you would expect workers exhibiting these traits to require minimal supervision. But, is this true? Is this how most current engineering managers lead? Looking at some current theories on leadership combined with personal anecdotes, this article will look at some common misconceptions about leading engineers.Engineers are different. There, it has been said. From cartoons to movies to television shows and more, there is a reason stereotypes of engineers resonate. Yes, you can find a lot of individual variation among specific engineers that you may know, but they do not detract from some of the overarching commonalities that engineers are known for.Clash of culturesWhat distinguishes a professional? Raelin (1985), in his book Clash of Cultures, describes professionals as having superior intellectual training, maintaining their own standards of excellence and being supported by associations that maintain the quality of the profession. Engineers obviously fit all three of these ideas: an engineering degree is still regarded as one of the most difficult to attain. Standards are maintained both through peer review and in academia, through ABET accreditation and through professional associations like SME to promote the discipline and foster technical growth.Six characteristics representing professional status are described by Raelin (1985, p. 9):• Expertise — prolonged specialized training in a body of abstract knowledge.• Autonomy — the freedom to choose the means to solve a problem.• Commitment — primary interest is in pursuing the practice of one’s own chosen specialty.• Identification — identifying with the profession and with fellow professionals, both through formal associations and through peers external to the organization.• Ethics — providing service without concern for oneself or without becoming emotionally involved with the client.• Standards — committed to help in policing the conduct of fellow professionals."

Jan 1, 2018

By S. Dixon

The Peñasquito deposit in Mexico is one of the largest grassroots precious metal discoveries developed in recent years. Goldcorp Inc. acquired the gold-silver-lead-zinc property with the acquisition of Glamis in 2006. The Peñasquito property is located in the northeast corner of Zacatecas state, Mexico. The elevation of the property averages 1,900 m (6,200 ft) above sea level. It is situated in a wide east-west valley. Minera Peñasquito, S.A. de C.V., wholly owned subsidiary of Goldcorp Inc. has operated twin SABC-flotation lines since June 2010. Each line is capable of processing 50,000 tonne per day using a SABC grinding circuit followed by flotation to produce lead and zinc concentrates. The addition of a Polysius Polycom HPGR 2417 with dual 3,000 kW drives to reduce the crushed pebbles produced by the SAG mills was planned to increase production to 65,000 tonne per day per line. The production has been limited to 100,000 tpd due to operating issues in the grinding circuit. The operating issue has been the amount of pebbles generated by the SAG mills is significantly less than originally planned. This paper describes the planned installation of a crushed augmented feed to maximize utilization of the HPGR capacity. This coupled with full utilization of the SAG and Ball Mill available energies there is the potential to achieve a 50% total capacity increase from the grinding circuit.

Jan 1, 2012

By D. R. Wilburn

This summary of international mineral exploration activities for the year 2007 draws upon available information from industry, literature and U.S. Geological Survey (USGS) specialists. The summary provides data on exploration budgets by region and mineral commodity, identifies significant mineral discoveries and areas of mineral exploration, discusses government programs affecting the mineral exploration industry and presents analysis of the mineral industry based upon these data. Two types of information are reported and analyzed in this annual review of international exploration for the year 2007: 1) budgetary statistics expressed in U.S. nominal dollars provided by Metals Economics Group (MEG) of Halifax, Nova Scotia, and 2) information on regional and site-specific exploration activities that took place in 2007 and compiled by the USGS.

Jan 1, 2008

By Herman Gill, Dani Delaloye, Madelyn Rubin

The Foothill Tunnel is a remote 16-mile-long untreated drinking water tunnel in the San Francisco Public Utilities Commission’s (SFPUC) Hetch Hetchy Water and Power (HHWP) system. The 14-foot horseshoe shaped tunnel, constructed in the 1920s, is lined intermittently with concrete over approximately 53% of its length, with the balance being unlined rock. After detailed planning, in January of 2020 Mott MacDonald and HHWP staff completed a visual inspection and documented the condition of the tunnel with electronic data collection, photographs, and video with audio recording. The result was a comprehensive condition assessment that identified key areas of the tunnel to monitor.

Jun 13, 2021

By Jerry G. Rose

This paper addresses a process by which both the disposal and utilization of preparation plant coal refuse can be simultaneously accomplished with an attendant benefit of significantly increasing the percentage of extractable coal. A laboratory study is described in which coal refuse was successfully utilized as an aggregate in low cost concrete (coal-crete) mixes under simulated mine environments. Coal-crete specimens with compressive strengths of 14,000 kPa (2,000 psi) were consistently produced. A construction sequence and mining plans are offered for placing the low-cost coal-crete pillars in accessible areas in room-and-pillar mines so that a Substantial portion of the remaining coal pillars could be extracted. The method could be applicable in areas where mining procedures do not permit pillar extraction because of possible surface subsidence, roof problems, or other conditions. Possible structural difficulties with the coal-crete pillar support system are discussed. Using a discounted cash flow analysis, the economic feasibility of mining with the coal-crete pillar system is investigated. The cost of the coal-crete pillar system is compared to the cost of a mining system which would have less coal recovery. The analysis indicates that although at higher coal selling prices the coal-crete mining system may yield an acceptable rate of return, the alternative of leaving the coal pillars is more attractive at the present time.

Jan 1, 1977

By Steve Karl

Introduction Technological advances in mining and processing have started the domestic mining and mineral industry back on the road to competitiveness. A massive trade deficit, however, coupled with foreign competition and more government regulations threaten to slow this recovery. So say mining executives in Colorado and Minnesota. These industry decision-makers agree that their challenge is to continue the technical innovations of their industry. The journey back to profitability and competitiveness means improved products, they say, but at lower production costs. At separate gatherings in January, officials in both states said that they believe their industry is up to the challenge. Worker attitudes will have to change, they understand, and management styles will need adjusting. And, while they are willing to make the necessary adjustments, they are going to need some help. Or, at the very least, less interference from the federal government. Foreign markets, all but closed to American products, must be opened up on a reciprocal basis. The nation's $170 billion trade deficit is a priority item that can no longer be ignored. And, closer to home, federal and state environmental regulations concerning mineral production must themselves be monitored. 1987 in Colorado While Colorado's service and trade industries should have a fairly good 1987, the state's mining industry can expect another disappointing year filled with layoffs and declining profits, according to the Colorado Mining Association. Once again, foreign competition and government regulations are seen as the main reasons for the mineral industry's woes. And state mining officials see little relief until at least the 1990s. Speaking to a group of reporters at its annual press briefing in Denver, CMA officials said Colorado's overall nonfuel mineral production is expected to fluctuate between a flat and declining growth trend. Mining employment in the state is expected to fall by 400 people this year, to about 7500. In 1981, Colorado mines employed as many as 17,800 people. During the 1970s and early 1980s, the employment rate in mining in the state grew by 12% annually, according to a summary report compiled by the CMA. Last year, worldwide molybdenum production continued to exceed demand US moly producers, therefore, were forced to continue their familiar path of production cutbacks, as well as workforce reductions. And some molybdenum mines in the US closed altogether, the CMA report stated. In Colorado, Amax's Climax and Henderson molybdenum mines operated at about 50% of combined capacity. Climax employed about 200 people at the beginning of 1987, producing about 5.4 kt/d (6000 stpd) or ore. Henderson employed about 700 people, producing about 24.5 kt/d (27,000 stpd) of ore. In coal, the number of licensed coal mines in Colorado has declined from 124 in 1960 to 43 in 1986, of which only 23 were producing. Once a stable source of employment and production, the metallurgical coal industry in the state has only one surviving metallurgical coal mine, Mid-Continent's Orchard Valley mine. High production and transportation costs, weakness in the US steel industry, and a weak market position have contributed to the decline, the summary reported. "The challenge facing most mining companies in the United States today is to survive," said

Jan 3, 1987

The recovery of coarse particles has been widely recognized as being limited by the froth in mineral flotation. In response to this, FLSmidth has developed the Selective Froth Recovery (SFR) System. The SFR System is an extraction system that vacuums material from select locations within the froth and separates the recovered solids into coarse and fine fractions with a hydrocyclone. The SFR is supported and located by an actuator connected to the cell?s level control so that the suction heads are maintained at an offset distance from the slurry level. This can increase the recovery of coarse material by selectively recovering the material close to the bottom of the froth where drop-back of coarse particles typically takes place. The SFR System has been tested at two separate mine sites, and has been proven to increase the copper recovery in both trials. TEST RESULTS The SFR prototype was installed in a copper recovery circuit, on a rougher scavenger 160-m3 flotation cell. Figure 1, shows the SFR System, which consists of four conical skimmer heads that are plumbed together and share a common actuator. The SFR System, raised out of the froth prior to a test, can be seen in Figure 2. The SFR prototype was able to maintain a variety of set points using the level control to determine the slurry height.

Feb 27, 2013

By J. C. Cawley

This National Institute for Occupational Safety and Health study was conducted to focus future research on the most significant electrical problems in the mining industry. Data from 1,926 mine electrical accidents (including 75 fatalities) that occurred between 1990 and 1999 were studied. Coal and metal-nonmetal operator- and contractor-reported data are presented. All data used in this analysis were MSHA closeout data, except 1999, which were preliminary data. Electricity was the 4th leading cause of death reported in mining despite being the 14th leading cause of injuries. Electrical shock, not electrical burns, caused 70 of the 75 electrical fatalities reported. About one-half of mine electrical accidents and fatalities occurred during electrical maintenance. Suggestions for mitigating electrical hazards are proposed.

Jan 1, 2002

By E. F. Perry

Simulated weathering tests are used to estimate leachate quality of rocks containing varying concentrations of pyrite and carbonates at mines. This study reports the inter-laboratory results from the third phase of a three part effort to develop a standardized, reproducible simulated weathering test for coal mine overburden. Initial development in phases I and II included evaluation of different test apparatus, effects of particle size, gas composition, leaching frequency and other variables. Phases I and II results are reported by Brady et al (2004) and Hornberger et al (2004). In Phase III, a standard column leaching protocol was used by eight commercial, government and university labs to test five rocks of different lithology, pyrite and carbonate content. Ideally, a simulated weathering test should have three characteristics to be useful in mine drainage studies. These include the capability to produce repeatable results, accuracy in simulating field conditions, and a concise test protocol of weeks to a few months. Simulated weathering, or kinetic testing, produces an effluent that simulates mine drainage composition. This can be beneficial for predicting mine drainage quality in comparison to whole rock analyses. The effluent may be tested for the same water quality parameters as the drainage produced during mining, such as pH, acidity, alkalinity, sulfate, iron, manganese, aluminum or other parameters. If the physical, chemical, and biological conditions of the kinetic tests are representative of those found in the mine environment, leachate composition may be used to estimate the water quality parameter concentrations produced during mining. Hornberger and Brady (1998), and Geidel et al (2000), compiled comprehensive chapters on kinetic tests for mine drainage prediction. These sources included: (1) a chronology and synopsis of scientific literature on these kinetic tests as they have developed over approximately 50 years, (2) discussion of the influence of physical, chemical, and biological processes, and (3) general guidelines for test procedures, data interpretation, and recommendations for further research to develop standard methods. Most of the kinetic test methods in use today were substantially developed and applied more than 40 or 50 years ago, including leaching columns (Braley, 1949), humidity cells (Hanna and Brant, 1962), Soxhlet reactors (Pedro, 1961), and field scale tests (Glover and Kenyon, 1962). . The most commonly used kinetic tests for mine drainage prediction are leaching columns and humidity cells. While these kinetic test methods and others have been used in hundreds of mine drainage studies as shown in Hornberger and Brady (1998) and other references (e.g. Sorini, 1997), they have rarely been used in coal mine permitting by either regulatory agencies or the mining industry. The major impediment to routine use of kinetic test methods is the variation in the design and operation of these kinetic tests, and lack of a standardized, accepted (e.g. by EPA or ASTM) test method. Interpretation and data comparison from different studies can be difficult. METHODS Five rocks of varying lithology, mineralogy, carbonate and pyrite content were collected in bulk from Pennsylvanian age coal measures in Pennsylvania, West Virginia and Indiana. Mineralogical composition of the samples was characterized in detail by Hammerstrom et al (in press) using optical, X-ray diffraction, scanning electron microscopy and electron microprobe methods. Acid Base Accounting (Sobek et al 1978; Skousen et al, 1997, Kania, 1998) was performed on bulk samples for % total sulfur (%S) as a measure of pyrite and potential acid generation, and Neutralization Potential (NP) as an index of carbonate content. A modified NP test, following the procedure suggested by Skousen et al, 1997 was also conducted for four of the five samples. The modified NP test accounts for initial acid consumption, Fe oxidation, and subsequent Fe hydrolysis by the iron carbonate mineral siderite. Siderite, even though it is a carbonate gives zero net alkalinity. Lithologic, Acid Base Accounting and general mineral composition data for the five rocks are shown in table 1 (see Appendix). Whole rock elemental analysis for major and trace elements was conducted on a sub-sample of each rock type. One laboratory was contracted to mix, homogenize, and size the bulk samples and prepare sub-samples for all the participating labs. Splits were sent to eight labs, and each organization was provided a detailed set of instructions for column assembly and leaching procedure. The protocol was developed based on results from phases 1and 2 of this study. The principal features of the column assembly are: a specific composition of particle size by weight (table 2), weekly leaching cycles, and incubation under humidified air containing 10% CO2. Figure 1 shows the column construction. [ ] The particle size distribution of the crushed rock sample is largely an artifact of the crushing process, rather than a natural systems process like the particle size distribution of a soil or an unconsolidated sedimentary deposit. The standardized particle size distribution, shown in Table 2, promotes operational consistency of the weathering test procedure, and provides better control over reaction kinetics. Large amounts of fine particles within specific zones of the leaching columns have been found to impede uniform fluid flow and/or gas flow (Hornberger and Brady, 1998). Regarding reaction kinetics, the importance of surface area to volume ratios has been described in a previous phase of this study (Brady et al., 2004), and significant differences in crushed particle size distributions and effective surface areas were found among the lithologic units. Standardizing the particle

Jan 1, 2009

By G. P. Tomaino

Talc is a layered, hydrous magnesium silicate mineral. It has a soft, soapy feel and typically a smooth texture. Talc is also known for its insulation, heat resistance, chemical stability, oil absorption and strong covering quality. Talc, Mg3Si4O10 (OH)2, has a theoretical chemical composition of MgO at 31.7 percent, SiO2 at 63.5 percent, and H2O at 4.8 percent. However, talc’s chemical and mineralogical composition can vary depending on its geological history/parent rock association. These mineral associations and variable levels are usually chlorite and carbonates (magnesite, calcite and dolomite). Two key elemental substitutions that can occur in the talc crystal structure are iron for magnesium and fluorine for hydroxyl. These compositional differences may limit or enhance the talc’s usage in specific market niches. The United States remains self-sufficient in most grades of talc.

Jan 1, 2005

By Minggao Qian, Jialin Xu, Wenqi Lai, Weibin Zhu

This article advances an innovative concept explaining the connotation and technical system of "green" mining. The important theoretical foundations of "green" mining are the distribution and development characteristics of joints, fractures and bed separations, and the seepage and flow behavior of methane and water in the broken rock strata following coal mining. The main focus of "green" mining technique includes water-preservation in mining areas, coal mining under buildings with bed separation spaces grouted to retard surface subsidence, partial extraction and backfill mining, simultaneous extraction of coal and coal-bed methane, coal roadway supporting and underground disposal of rock waste, underground coal gasification, etc.

Jan 1, 2005

The efficiency of a coal mine ventilation circuit is usually measured by the ratio of the sum of the face quantities divided by the quantity supplied by the fans. Other types of systems use the power delivered to the power supplied to measure efficiency. This paper argues the case that the current measure may be inadequate and that by also looking at the power ratio, ventilation may be better managed. Problems associated with failing ventilation systems would be highlighted earlier so that crisis management could be avoided.

Jan 1, 2009

By M. J. Sapko

The design requirements for mine ventilation seals have undergone a radical transformation in recent years, and these revisions have greatly increased the cost of the seal designs and their construction. For example, in the past two years, new federal regulations have increased the minimum design requirement to withstand explosion pressures from 138 kPa (20 psig) to 345 kPa (50 psig) or 827 kPa (120 psig), depending on whether the sealed mine volume is monitored continuously or not. Moreover, there is still a possibility that under certain conditions (such as detonations) even higher pressure requirements may be necessary. The ability of a monolithic, stand-alone mine seal to reliably withstand the full range of current and future pressures is becoming increasingly difficult and expensive. In an effort to develop a practical alternative, the West Virginia Office of Miners? Health Safety and Training (WVOMHST) and the National Institute for Occupational Safety and Health (NIOSH) have collaborated in a research effort to develop a practical, economic, and safe mine sealing technique that can enable mines seals to meet the full range of new explosion pressure design requirements. The basic idea is to use a barrier of common mine gob and rubble in combination with a conventional mine seal so that the pressure resulting from a gas explosion is reflected, absorbed, and attenuated, so the pressure on the ventilation seal is reduced. This paper discusses some concepts and preliminary test results for a ?Passive Mine Blast Attenuator? (PMBA) that can provide a useful alternative to increasingly larger and stronger stand-alone mine seals. Numerical models and full-scale experiments conducted at the NIOSH Lake Lynn Laboratory Experimental Mine (LLEM) show that the use of a PMBA can signficantly reduce the blast pressure and impulse on conventional ventilation seals. DISCLAIMER: This information is distributed solely for the purpose of pre-dissemination peer review under applicable information quality guidelines. It has not been formally disseminated by the National Institute for Occupational Safety and Health (NIOSH). It does not represent and should not be construed to represent any agency determination or policy.

Jan 1, 2009

By K. G. Correia

The replacement of the dry mining method at Guajú Mine by the dredging method was a means of ensuring its feasibility by maintaining the production levels. Dry mining increased the cost of the deposit with the distance and presented a limited mined volume, while the decreasing grade of the deposit led to a production drop. With the new methodology, there was a considerable reduction in unit cost of the mine and an increase in the production level of a deposit with low heavy minerals content.

Jan 1, 2005

By T. McCallum, B. Wellbrook, T. Ratnaweera

"Pressure oxidation (POX) leaching of copper (Cu) concentrates continues to be explored as the next generation of leaching processes to sustain existing Cu Solvent Extraction/Electrowinning (SX/EW) operations as easily leachable ores become scarcer. The leach liquors produced, particularly from total pressure oxidation, contain elevated concentrations of Cu and acid, making solvent extraction processing more challenging. Addressing this challenge will require innovative processing techniques and continuous improvement to be successful in achieving satisfactory physical and metallurgical performance in the SX plant. The ability to efficiently process these feeds via SX plays an important role in the continued success of POX technology. Characteristics of a concentrate leach pregnant leach solution (PLS) are discussed, along with best practices that should be considered for processing by SX. Findings discussed in this paper will shed light on ways to reduce process risk and bring additional improvement to the industry with minimal change to the SX plant. INTRODUCTION Hydrometallurgical processing of copper concentrates continues to gain interest from the industry. Treatment by smelting and electro refining continues to dominate, accounting for ~80% of global copper production, but is faced with its own challenges. The case for increased hydrometallurgical processing has been discussed as high grade ore deposits are diminishing, capital costs for smelting are high, and the environmental aspects present challenges for sustainable practice. The hydromet copper industry is also turning a corner as easily leachable copper oxide deposits are depleting, and excess SX-EW capacity exists in certain regions. The transition to hydrometallurgical processing of copper concentrates will take a bit of effort as leaching of copper sulfides is difficult and byproduct credits must be considered. Various POX leaching processes have been studied and proposed to the industry as viable practices. Of the many processes, commercial use has been practiced with the Mount Gordon Copper Process, total and partial pressure oxidation by Freeport-McMoRan at their Bagdad and Morenci, Arizona complexes, the Sepon Copper Process in Laos, and the bioleach BIOCOP™ process in Chile. The experience related to the current paper centers around Cu concentrate leaching performed by Freeport McMoRan (known as CLP). The patented process was developed as an alternative to smelting and refining in 1998. The technology applies to both molybdenum and copper concentrates. Commercial CLP research and development began with Phelps Dodge in 1999 to 2001. The first commercial plant was commissioned in 2003 at the Bagdad operation. Initial operating conditions of the Bagdad CLP were under high temperature conditions before converting in 2005 to medium temperature conditions. In 2007-2009, the CLP technology was commissioned in Morenci, operating medium temperature and high temperature lite conditions until shutting down due to the economic downturn. The Morenci CLP restarted operation in April 2015 due to acid production benefits for the Morenci leach operations. Renewed solvent extraction test work initiated well before restart and has been an ongoing focus for further SX plant optimization."

Jan 1, 2018

By Harry C. Verakis

A conveyor belt fire presents a serious safety hazard in an underground coal mine. A review of underground coal mine fires involving conveyor belts which occurred within the past two decades was made by the Mine Safety and Health Administration (MSHA). Belt fires that occurred during this period of time have burned as much as 610 m of belting before the fire was extinguished. In an effort to improve safety, MSHA has undertaken a program to study the flammability of conveyor belts in large and small-scale tests. Large-scale tests were conducted on various types of mine conveyor belts using a range of airflows from 0 to 6.1 m/s. Data were obtained to evaluate the effect of airflow on flame spread and other combustion properties of the conveyor belts. The results from the large-scale tests were utilized to develop a new MSHA laboratory-scale test for evaluating the flame resistance of conveyor belts used in underground coal mines. Agreement of results based on a pass/fail criterion was obtained between the large-scale test and new laboratory-scale test. A summary of test results from the new test is shown for different conveyor belt formulations. MSHA also performed a series of fire tests with wood samples using the new laboratory-scale tunnel apparatus. These tests were conducted at airflows from 0.5 to 5.1 m/s to determine the effect on flame spread and burn damage. The data were also analyzed to determine if wood could be used as a calibration standard. The results indicate the flame spread behavior of the wood follows trends similar to results obtained with conveyor belts. The data did not show the precision needed to use wood as a calibration standard. MSHA intends through rulemaking procedures to replace the "20g" conveyor belt acceptance test with the new laboratory-scale test. MSHA also plans to implement a voluntary acceptance program. Under this program, a. conveyor belt manufacturer could obtain an MSHA acceptance for belts meeting the new test criterion. A belt accepted under the voluntary program could then be used underground. Use of conveyor belts meeting the new test would limit the propagation of fire and reduce the probability of igniting other combustibles in the belt entry.

Jan 1, 1991

By T. P. Dolley

Domestic production of industrial sand and gravel in 2004 was about 29 Mt (32 million st), a 6-percent increase from 2003. This increase was buoyed by a robust construction sector of the U.S. economy. U.S. apparent consumption was 30.3 Mt (33.3 million st) in 2004, an in-crease from 2003. Imports were about 490 kt (540,000 st) in 2004, an 11-percent increase from 2003. Imports of silica are generally of two types. They include small-quantity shipments of very-high-purity silica or a few large shipments of lower grade silica that were shipped only under special circumstances (such as very low freight rates).

Jan 1, 2005

By A. Martin

Detailed porewater profiles of arsenic and other interstitial metabolites were used to assess the relationship between trophic status, sediment redox chemistry and arsenic mobility in a historic tailings repository. The results demonstrate that in eutrophied systems which exhibit greater sediment oxygen demand, the depth of the Fe(III) redoxcline, and hence depth of arsenic release, may lie in close proximity to the sediment-water interface. As a result, arsenic scavenging mechanisms are significantly reduced, resulting in a greater flux of arsenic to the water column. Such considerations have important implications for the management of pit lakes, tailings ponds and contaminated lakes which may consider: 1) the discharge of sewage to the water body during operations; and/or 2) post-closure fertilization as a remediation strategy.

Jan 1, 2002