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By Leroy O. Gieseke, John B. Begley

If the particles produced and transported in your plant are as chunky as rough crushed coal or as fine as cement, there are probably several conveying and elevating systems suitable for your operation. The problem is deciding which is most suitable in terms of space, energy cost, allowable product degradation, productivity, and angle of movement. Each conveying and elevating method should be fit to the products it can handle successfully, to desirable economics, and to suitability for the overall operating scheme. Matching along these lines will result in selection guides that should be useful for many years, although design evolution is expected for the materials handling equipment involved. There are several common bulk material handling devices for moving solids: belt conveyors, screw conveyors, bucket elevators, chain conveyors, and en-masse conveyors/ elevators. En-masse conveyor/elevator equipment is commonly compared and contrasted with the other devices. These comparisons often involve en-masse versus all others. En-masse has developed into a versatile technique because the machines can move materials horizontally, vertically, or in closed circuit. Such machines unload, elevate, store, reclaim, and distribute materials-all without ancillary equipment except for chutes and gates. Bulk material can be carried, as with a belt conveyor or bucket elevator; pushed, as with a flight or screw conveyor; bumped, as with a vibrating conveyor; blown, as with a pneumatic conveyor; fluidized, as with a dense phase pneumatic conveyor; or by applying internal forces to it to induce "en-masse" flow. Conveying en-masse is analogous to the flow of a fluid in an open channel, and elevating en-masse is analogous to the flow of fluid in a pipe. In the case of a fluid, flow is caused by a differential in pressure. With en-masse movement, flow results from resistance to shear induced by the movement of chain and flights.

Jan 12, 1980

By M. Fornaro, E. Lovera, I. Sacerdote

"The analysis of the mining laws of the different European States reveals big diversities and, often, remarkable complications. This is due, firstly, to the history and culture of different States and, secondly, to the assets of natural resources. It is therefore difficult to find a common line between the text of the various mining laws, but it can be observed that nearly all the States distinguish the mining resources in two categories: the materials whose importance is considered strategic belong to the first category; in the second category the materials whose importance (above all economic) is mainly local are comprised. The paper, analysing in detail the planning procedures in some important Countries, investigates the possible perspectives of arranging some guide-lines for a correct approach to the sustainable exploitation planning/programming of mineral resources (with particular reference to "quarry" materials, such as aggregates), considering both the environment protection and the production needs."

Jan 1, 2005

By E. Barnard

Tarkwa Gold Mine is located in south western Ghana, and is majority owned and operated by Gold Fields Ghana Limited. The Tarkwa CIL plant utilises a single low aspect 27' x 42' 14MW dual pinion drive SAG mill designed to mill 4.2mtpa (525tph). The plant was designed and constructed in a manner that would facilitate its expansion to 8.4mtpa (1,050tph). Since the plant was commissioned in September 2004, the SAG mill has consistently exceeded its nameplate throughput by up to 25%. Through the installation of a 26? x 36? 14MW dual pinion drive Ball mill along with additional upstream and downstream equipment, the plant will now be expanded to 12mtpa (1,500tph). This paper endeavours to outline the process and engineering design challenges along with the future construction challenges of the expansion in the midst of a fully operational plant.

Jan 1, 2008

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 P. Roghanchi

The Rock Mass Rating (RMR) system is an internationally recognized classification system that can be used for preliminary ground control design. The RMR system assigns quantifiable values to predefined classified parameters of a rock mass. Assigning a single value rather than a range to each parameter is a source of uncertainty, however, assigning a single value can be useful for further studies related to RMR ratings. Fuzzy systems are being used in an increasing number of application areas where linguistic rules are used to describe the system. In past decades, fuzzy systems have been successfully used in geotechnical and mining engineering problems to cope with uncertain data as well as with vagueness. In such scenarios, linguistic rules and rating-based classification of rock masses have always been questionable. The objective of this study is to apply fuzzy set theory to the RMR classification system. The fuzzy system proposed in this study is specified for weak and very weak rock masses (RMR<50) based on data collected from Nevada gold mines where 68% of RMR data is in the ?poor? and ?very poor? classes. In this approach, a Mamdani fuzzy algorithm has been constructed to evaluate basic RMR89 rating for the ?wet condition.? Comparison between basic RMR89 and fuzzy RMR shows the fuzzy RMR system is valid to predict RMR from the discrete rating of each input parameter.

Feb 27, 2013

By D. A. Kramer

Ammonia, the principal source of fixed nitrogen, was produced by 16 companies at 31 plants in the United States during 2003. Fifty-five percent of U.S. ammonia production capacity was centered in Louisiana, Oklahoma and Texas because of their large reserves of natural gas, the dominant domestic feedstock. Production In 2004, U.S. producers operated at about 59 percent of their rated capacity. The United States remained the world’s second largest ammonia producer and consumer following China. Urea, ammonium nitrate, ammonium phosphates, nitric acid and ammonium sulfate were the major derivatives of ammonia in the United States, in descending order of importance.

Jan 1, 2005

By R. Owen

An overwhelming amount of data can be collected around the blasting process. This information can include blasting product, pattern design, blast results, and routing. It is only when these data sources are combined in a concise and accurate form that they are of real use in determining the safety and economic implications of each parameter. Two key technological aspects of blasting are drill fleet management (DFM) and size fraction analysis (SFA). As these technologies are integrated as near-real-time measurement and QA/QC tools, the resolution and realization of blasting parameters is significantly increased. The Freeport-McMoRan Morenci Mine has integrated these technologies along with other existing database structures to create a responsive and sustainable tool for reconciliation, forecasting and parameter matching. This ability allows Morenci to customize blast patterns to meet the criteria most critical to each shot, whether the impact is influenced by routing and recovery, equipment maintenance or safety concerns. DRILL FLEET MANAGEMENT The current DFM was stabilized in its current form in mid 2009 after database and hardware maintenance met stability requirements. Currently the system is running on 12 production drills. The system receives pattern layout and is capable of semi-autonomous drilling over the entire district. The fleet is capable of drilling in highly altered intrusive units and hard granites through the range of sedimentary and volcanic units that comprise the Morenci District. The system brings a level of accuracy not achievable with earlier paint-marked patterns. Depth control is also increased along with more consistent wear and maintenance patterns due to semi-autonomous drilling.

Jan 1, 2011

By K. T. Lee, M. U. Ozbay, M. D. G. Salamon

Design of yielding chain pillars in deep coal longwall mines is not as well established due mainly to the lack of knowledge in the behavior of coal pillars in the post-peak regime. This study shows that two of the existing strain-softening models have the potential to provide an improved understanding of yielding and, therefore, could form a rational basis for yielding coal pillar design. Analytic and numerical modelings of a row of chain pillars in a two-entry development at a depth of nearly 700 m suggests that yielding pillars are feasible, but the design is very sensitive to the coal’s post-peak behavior.

Jan 1, 2001

By D. Malhotra

The economic viability of a mining operation depends on continuous process improvements. Diagnostic testing and plant auditing are essential useful tools to improve performance and reduce operating costs. Case studies are discussed to illustrate benefits of these techniques to improve plant efficiency.

Jan 1, 2001

By Jerry D. Pfarr, Gary Loving, David G. McDowall

Introduction The Open Cut is one of the best known landmarks in South Dakota&apos;s Black Hills. It is the site of the original discovery of the Homestake ledge, or lead, in 1876. Between that time and 1945, the Open Cut produced about 36 Mt (40 million st) of gold ore, the majority of which was drawn from underground operations. In 1981, a drilling program was undertaken to determine how much ore the old timers left behind. The results of this study indicated an ore body made up of broken and caved ground, and solid blocks in pillars and crowns. Reserves are estimated at about 6.8 Mt (7.5 million st) of gold ore, grading at 4 g/t (0.116 oz per st). Total tonnage of material, including waste, is estimated at more than 60 Mt (66 million st). The mining plan is a combination open pit and underground operations. The bottom 78 m (255 ft) will be mined by the underground panel-caving method. The open pit was designed as a contractor operation up to the expected milling stage or the point of sustained ore production. The pit was designed to avoid disturbance of US Highway 85 near Lead, SD. In 1983, the South Dakota Board of Minerals and Environment approved the mining and reclamation plan. Test mining began in the north of the Open Cut using open pit methods. Upon completion of the test pit, Homestake Mining&apos;s board of directors approved an expanded mine plan. They committed an additional $27.1 million to prestrip the ore body that is expected to yield 23 t (750,000 oz) of gold. Project life is estimated at 15 years. Exploration program In early 1980, 10 diamond drill holes were drilled from the surface in the vicinity of the east side of the Open Cut. The reserves delineated by that drilling were too small to justify their development. A continuation of this study attempted to quantify the tonnage remaining in the bottom of the Open Cut in the crushed remnants of old stope-and-crown pillars. This was originally thought to be minable by improved drawhole methods using highly mechanized procedures. By early 1981, however, thinking turned toward surface mining methods. The effort to quantify minable ore tonnage ultimately involved drilling 385 reverse-circulation rotary drill holes. These were drilled from August 1981 to March 1983 in several groups of holes. Each group, or drilling program, had a slightly different immediate goal or purpose. The holes were drilled to various depths using truck-mounted Drill Tech 40 drills with doublewalled 76 mm (3 in.) pipe with 134 mm (5.25 in.) three-cone rotary bits. Initially, wet drilling was employed using a cyclone collection system. Since cutting recovery was not as high as it should have been, all subsequent drilling was dry. The cuttings were collected and deposited by way of a cyclone into the sample bag. The pickup for the cuttings using the three-cone bits was about 50 mm (2 in.) above the bottom of the hole. Overall cutting recovery for the drilling program was estimated at 75%. Samples were taken on each 1.5 m (5 ft) of hole drilled. Each sample was split into three 48-g (1-2/3 assay ton) charges. The accepted value was the straight average of the three charges. Because of the known "nugget effect," assays were nearly impossible to duplicate. Therefore, to validate the regular assays, a series of screen fire assays were run. Based on this method of assays and checks, it was concluded that 10% more ounces would be produced than indicated by mine grade. Test pit The northernmost portion of the mine as planned is known as the Terraville test pit. This portion was mined ahead of the rest and separately. It provided answers to questions on slope stability, reserves correlation, mill recoveries, mining costs, and surface mining conditions in various weather and ground conditions since many boulders and much timber were expected. The mining plan for this portion of the pit was successful. More ore was mined than was expected. Mixing of Open Cut ore with underground ore in the mill feed was successful. Impact on the surrounding communities

Jan 10, 1987

By Richard Addison

Introduction A staccato beat rings out across the Australian bush. The stamps are at work, slowly pulverizing gold-bearing ore for a patient, grizzled prospector. This is the State Battery near Marvel Loch, a tiny 22.6-t/d (25-stpd) capacity gold plant operated by the Western Australia government. It is one of many such plants located in remote corners of Western Australia and the Northern Territory. They are miniature custom mills operated for the benefit of prospectors and small mine operators. These state batteries are so called because they are operated by state governments and because the principal item of machinery in the plants are stamp batteries. They are fascinating anachronisms - throwbacks to times past, a century ago, to the hard and heady existence of gold rush days. Marvel Loch plant Typical of these operations, and perhaps a trifle more sophisticated than most, is the plant near Marvel Loch, a small center for mining and farming about 400 km (250 miles) due east of Perth. Marvel Loch. The name evokes visions of a pristine lake set in a cool, green, misty glen of the Scottish highlands. Actually, it is probably a hot, dry salt pan somewhere nearby. Just where the "loch" is difficult to discern. There is no sign of it on the map.

Jan 7, 1986

By Norbert Patzelt

Today the standard method for grinding ore is the wet process. This is mainly due to the fact that the downstream process is wet and considering the total circuit dry grinding is not the most eco­nomical method for the comminution of the ore. Presently in gold ore processing dry grinding is only applied in cases where whole ore roasting is required downstream from the milling plant or if organic carbonaceous matters and sulfides are the primary reasons for the refractoriness. Since dry ore can be fed to the roaster, a dry grinding mill can advantageously be applied. In cases where d1y grinding could be applied advantageously for instance if the ore contains water soluble material and wet grinding is not possible there are serious concerns prevailing within the industty against this technology due to several reasons. Besides of the aspects as discussed later in this paper, there are several plants in operation with little success and considerable operational prob­lems. The reasons for this often lie in the fact that the several com­ponents of such a plant are bought from independent sources with­out considering the system requirements. It is essential that the dry grinding plants require accurate design between drying, pneumatic transport and grinding of the ore in the circuit. This paper will address some of these concerns raised in the industiy from a mill vendor's view along with discussing the development for the selection and configuration of the various dty grinding circuits.

Jan 1, 1998

By D. Tao

The reverse cationic flotation is the most widely utilized flotation method for concentrating iron ore. In such a process quartz is often floated with ether amines (R?O(CH2)3?NH2) partially neutralized with acetic acid. However, it is well known that the cationic flotation of hematite does not work well if a certain amount of slime is present in ore slurry. To improve flotation performance, starch is often added to the flotation feed to selectively flocculate and depress the iron minerals and sodium silicate is used to disperse the silica gangue. The present study was conducted to improve the flotation efficiency of iron ore flotation by use of a new type of depressant for clay and iron minerals, i.e., GP clay binder. It is postulated that the GP clay binder minimizes slime adsorption on quartz and hematite particles by agglomerating clay particles to reduce their surface area. It is further theorized that the clay binder helps depress hematite flotation due to its unique chelating ability to iron ion. The results from this study clearly show that the GP clay binder is an effective depressant in iron ore flotation. When it was used as a substitute for corn starch, it increased concentrate grade from 60.78% to 67.67% and increased iron recovery from 71.96% to 72.90%. The combined use of GP clay binder with corn starch produced even better flotation performance, increasing concentrate grade to 67.08% and iron recovery to 78.15% simultaneously. Key words: Clay Binder; Depressant; Iron Ore; Flotation.

Jan 1, 2009

Due to deposit shape and dimension, narrow vein makes reserve modeling, geological control, stope design, drill and blasting more challenging. It needs high standard quality control as errors can be created easily, which could cause significant amount of ore loss or dilution. From channel sampling, survey control, reserve modeling, drill hole design, blasting, communication with underground drillers to backfill, this paper illustrated every step and its quality control of a long hole stope mining. Compared to other narrow vein mining methods (cut and fill and shrinkage), long hole stoping means higher production, less labor, and significantly lower cost when deposit characteristics allowed.

Jan 1, 2012

By Eung Ha Cho, Wendi Li

Wet desulfurization of Pittsburgh No. 8 coal and Illinois No. 6 coal were conducted with sodium hypochlorite in the laboratory. Pittsburgh No. 8 coal was leached by hypochlorite at high pHs in one step only. The hypochlorite concentration varied from 0.2 to 0.6 molar; sodium hydroxide concentration, from 0.2 to 0.8 molar; and temperature was at the levels of 80 and 90°C. The desulfurization of Illinois No. 6 coal was conducted in three consecutive steps of pretreatment in concentrated ammonia at room temperature, leaching with hypochlorite at room temperature and hydrolysis in sodium hydroxide solution at 90°C. In the leaching step, hypochlorite concentration varied from 0.2 to 0.8 molar. The desulfurization method of Pittsburgh No. 8 coal was found to reduce mainly pyritic sulfur. More than 70% of pyritic sulfur removal was achieved at the optimum conditions of 0.4 molar hypochlorite, 0.4 molar sodium hydroxide and 90°C. The desulfurization method for Illinois No. 6 coal was capable of reducing significant amounts of organic sulfur. The removal of organic sulfur achieved a 37.8% reduction at the optimum operation of leaching at 0.4 molar hypochlorite and room temperature followed by hydrolysis at 0.3 molar sodium hydroxide and 90°C. The chlorine content in the coal produced by the chlorination during leaching was kept below the threshold value of 0.3% at the optimum conditions of the desulfurization method for each coal.

Jan 1, 2005

By Byron T. Darnton

Gravity concentration at the Pegasus Gold/Montana Tunnels Mine, near Helena, MT, began in June 1991 when a coarse-gold-scalping circuit was commissioned. It processes a small portion of the ball-mill circulating load to reduce the risk of gold particles that are too coarse to float from advancing to flotation. The circuit also cost-effectively enhances overall gold recovery and payable gold ounces. This gravity circuit has been a major success. Since the circuit was commissioned, throughput has increased 30%. Several modifications and improvements were made during the last three years that allowed the gravity circuit to keep pace with the higher tonnage. The overall approach to gravity concentration at Montana Tunnels has always been to optimize production rather than maximize the gold ounces produced. This philosophy has now yielded a circuit expansion and relocation project. It will allow for an estimated 50% increase in gravity gold production, while significantly reducing the unit costs of production.

Jan 1, 1995

By D. W. Olson

The estimated value of natural gemstones produced from U.S. deposits during 2004 was $12.9 million. This was a 3-percent increase from that of 2003. U.S. gemstone production included agates, amber, beryl, coral, garnet, jade, jasper, opal, pearl, quartz, sapphire, shell, topaz, tourmaline, turquoise and other gem materials. More than 60 varieties of gemstones have been produced commercially from U.S. mines. But commercial mining of gemstones has never been extensively undertaken in the United States. Most of the deposits have been relatively small compared with other mining operations. In the United States, much of the current gemstone mining is conducted by collectors, gem clubs and hobbyists rather than business organizations.

Jan 1, 2005

By J. D. Potts

Installation of a horizontal air blocking shelf around the inside perimeter of the surface mining drill deck shroud is proposed to improve dust capture. Without the use of this shelf, a strong coanda effect airflow pattern results in dust leakage to the environment, especially at the shroud-to-ground interface. Bailing air exits the drill hole and travels along the drill steel to the underside of the table, where it fans out across the bottom of the drill table and continues down the sides of the shroud. This dust-laden bailing air then strikes the ground and fans out under gaps between the shroud and the ground. The proposed shelf prevents dust-laden air from striking the ground by redirecting it toward the center of the enclosure, where it can be captured. Laboratory tests demonstrated that the shelf reduced dust levels outside of the shroud by 81% under operating conditions consisting of a 1.9:1 collector to bailing airflow ratio, and a 5.1-cm (2-in) gap between the shroud and ground. The shelf was 15.2-cm (6-in) wide and located at a height equal to 2/3 the distance measured from the bottom of the shroud to the bottom of the drill table. The air blocking shelf was also capable of overcoming the adverse impact of vertical leakage along the shroud seams as long as this leakage was confined to an area below the horizontal plane defined by the shelf location.

Jan 1, 2009

By Andreas Prucker

The Greater Vancouver Water District (GVWD) is currently constructing the largest water treatment plant in Canada. As part of this treatment scheme twin water transfer tunnels are being constructed. Work commenced with the 11 m diameter, 185 m deep Seymour Shaft which was excavated by drill and blast using a mobile crane for hoisting. The Seymour Shaft is the first large diameter shaft that has been sunk in Western Canada in recent years. The operational aspects associated with the hoisting, the drill-blast-support cycle and water control in the shaft are the focus of this paper.

Jan 1, 2008

There is wide agreement that sustainability has multiple dimensions, but substantially less agreement on the appropriate balance among those dimensions. This occurs because peoples’ vision of a sustainable future, their goals for sustainable development, and their objectives for resource management are reflections of their values, culture, and life experiences. Even assuming agreement about the objectives of sustainable resource management can be reached; there is still a need for policy and decision makers to choose from among alternative courses of action, each of which has its own set of impacts and outcomes. One of several tools used to compare the consequences of alternative actions is the Computable General Equilibrium (CGE) model. This paper uses CGE modeling to demonstrate how policy decisions do not take place in a vacuum and how policy-makers must be aware that there are trade-offs involved in any decision.

Jan 1, 2005