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Rules of Thumb for Mining and processing

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    maintenance

    Mining Field: Mine Maintenance
    Area: General •The degree of maintenance enforcement at an operating mine should be just less than the point that disruptions to operations are at a level where additional maintenance costs equal the resulting profits from production. Source: David Chick •In a trackless mine operating round the clock, there should be 0.8 journeyman mechanic or electrician on the payroll for each major unit of mobile equipment in the underground fleet. Source: John Gilbert •Emergency repairs should not exceed 15% of the maintenance workload. Source: John Rushton •LHD units at a shallow mine with ramp entry should have a utilization of 5,000 - 6,000 hours per year. Source: Unknown •Captive LHD units should have a utilization of 3,500 - 4,500 hours per year. Source: Unknown •LHD units in production service should have a useful life of at least 12,000 hours, including one rebuild at 7,500 hours. A longer life can be presumed from LHD units at the high end of the market with on-board diagnostics. Source: John Gilbert •Underground haul trucks should have a useful life of 20,000 hours; more if they are electric (trolley system). Source: John Chadwick
    Area: Service •An efficient Maintenance Department should be able to install one dollar worth of parts and materials for less than one dollar of labor cost. Source: John Rushton •A servicing accuracy of 10% is a reasonable goal. In other words, no unit of equipment should receive the 250-hour service at more than 275 hours. Source: Larry Widdifield
    Area: Infrastructure •With ramp entry, a satellite shop is required when the mean mining depth reaches 200m below surface. A second one is required at a vertical depth of 400m. Source: Jack de la Vergne •With ramp and shaft entry, a main shop is required underground when the mean mining depth reaches 500m below surface. Source: Jack de la Vergne •A main shop facility underground should have the capacity to handle 10% of the underground fleet. Source: Keith Vaananen •Service shops for open pit mines should be designed with plenty of room between service bays for lay-down area. As a rule of thumb, the width of the lay-down between bays should be at least equal to the width of the box of a pit truck. Source: Cass Atkinson

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    Mining Field: Explosives and Drilling
    Area: Powder Consumption •Powder consumption for shaft sinking is 2.5 lb./short ton broken.
    Listed below is typical powder consumption in hard rock.
    Shaft Sinking – 2.5 Lb./short ton broken
    Drifting – 1.8 Lb./short ton broken
    Raising – 1.5 Lb./short ton broken
    Slashing – 0.8 Lb./short ton broken
    Shrink Stope – 0.5 Lb./short ton broken
    O/H Cut and Fill – 0.5 Lb./short ton broken
    Bulk Mining – 0.4 Lb./short ton broken
    Block Cave u/c – 0.1 Lb./short ton to be caved
    Open Pit Cut – 0.9 Lb./short ton broken
    Open Pit Bench – 0.6 Lb./short ton broken
    Source: Various
    Area: Explosive Choice •The strength of pure ammonium nitrate (AN) is only about one-third as great as that of an oxygen balanced mixture with fuel oil (ANFO). Source: Dr. Melvin Cook
    Area: Blasting Strength •Blasting strength is a direct function of density, other things being equal. Typical explosives for dry ground (ANFO) may have a blasthole density (specific gravity) of 0.8 to 1.3, while for wet ground (slurry or emulsion) it varies from 1.1 to 1.3. Developments in explosive technology make it possible to choose any density desired, within the given ranges. Source: Dr. Nenad Djordjevic
    Area: Spacing and Burden •For hard rock open pits or backfill rock quarries, the burden between rows can vary from 25 to 40 blasthole diameters. Spacing between holes in a row can vary between 25 and 80 blasthole diameters. Source: Dr. Nenad Djordjevic •To obtain optimum fragmentation and minimum overbreak for hard rock open pits or backfill rock quarries, the burden should be about one-third the depth of holes drilled in the bench. Source: Dr. Gary Hemphill •To obtain optimum fragmentation and minimum overbreak for stripping hard rock open pits or quarrying rock fill, the burden should be about 25 times the bench blasthole diameter for ANFO and about 30 times the blasthole diameter for high explosives. Source: Dr. Gary Hemphill •The burden required in an open pit operation is 25 times the hole diameter for hard rock, and the ratio is 30:1 and 35:1 for medium and soft rock, respectively. The spacing is 1 to 1.5 times the burden and the timing is a minimum of 5 ms (millisecond) per foot of burden. Source: John Bolger •The burden and spacing required in the permafrost zones of the Arctic is 10-15% less than normal. Source: Dr. Ken Watson •When "smooth wall" blasting techniques are employed underground, the accepted standard spacing between the trim (perimeter) holes is 15-16 times the hole diameter and the charge in perimeter holes is 1/3 that of the regular blastholes. The burden between breast holes and trim holes is 1.25 times the spacing between trim holes. Source: M. Sutherland
    Area: Collar Stemming •The depth of collar for a blasthole in an open pit or quarry is 0.7 times the burden. Source: John Bolger •The depth of collar stemming is 20-30 times the borehole diameter. Source: Dr. Nenad Djordjevic •For open pits or back-fill rock quarries, pea gravel of a size equal to 1/17 the diameter of the blasthole should be employed for collar stemming (i.e. ½ inch pea gravel for an 8½-inch diameter hole). Source: Dr. Gary Hemphill
    Area: Relief Holes •Using a single relief hole in the burn cut, the length of round that can be pulled in a lateral heading is 3 feet for each inch diameter of the relief hole. For example, a 24-foot round can be pulled with an 8-inch diameter relief hole. Source: Karl-Fredrik Lautman •It has been found that a relief hole of 250 mm (10 inches) will provide excellent results for drift rounds up to about 9.1m (30 feet) in length. Source: Bob Dengler
    Area: Blastholes •The cost of drilling blastholes underground is about four times the cost of loading and blasting them with ANFO. Present practice is usually based on the historical use of high explosives where the costs were about equal. An opportunity exists for savings in cost and time for lateral headings greater than 12 feet by 12 feet in cross-section by drilling the blastholes to a slightly larger diameter than is customary. Source: Jack de la Vergne •The "subdrill" (over-drill) for blastholes in open pits is 0.3 times the burden in hard rock and 0.2 times the burden in medium/soft rock. Source: John Bolger •"Sub-grade" (over-drill) is in the order of 8 to 12 blasthole diameters. Source: Dr. Nenad Djordjevic
    Area: Ground Vibration •The ground vibration produced by the first delay in a burn cut round is up to five times higher than that generated by subsequent delays well away from the cut. Source: Tim Hagan
    Area: Crater Blasting •Crater blasting will be initiated if the charge acts as a sphere, which in turn requires the length of a decked charge in the blasthole to be no more than six times its diameter. Source: Mining Congress Journal
    Area: Labor Cost •The labor cost for secondary blasting can be expressed as a percentage of the labor cost for primary mucking. For Sub-Level Cave and Crater Blasthole stoping, it is around 30%; for Sub-Level Retreat it is closer to 10%. Source: Geoff Fong
    Area: Drilling •Percussion drilling is required for drilling blastholes in rocks with a hardness of 4 or greater on the Mohs' scale (Refer to Chapter 1). These are mainly the volcanic rocks. Rotary drilling is satisfactory for softer rocks, mainly sedimentary. Source: Dr. Gary Hemphill

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    Given: 8
    Mining Field: Mine Dewatering
    Area: Water Balance •The average consumption of service water for an underground mine is estimated at 30 US gallons per ton of ore mined per day. The peak consumption (for which the water supply piping is designed) can be estimated at 100 USGPM per ton of ore mined per day. Source: Andy Pitz •Ore hoisted from an underground hard rock mine has moisture ******* of approximately 3%. Source: Larry Cooper •A water fountain left running underground wastes 1,100 USGPD. Source: Jack de la Vergne •A diesel engine produces 1.2 litres (or gallons) of moisture for each litre (or gallon) of fuel consumed. Source: John Marks •In the hard rock mines of the Canadian Shield, ground water is seldom encountered by mine development below 450m (1,500 feet). This may be because the increased ground stress at depth tends to close the joints and fractures that normally conduct water. Source: Jim Redpath
    Area: Layout •The main pump station underground must have sufficient excavations beneath it to protect from the longest power failure. The suggested minimum capacity of the excavations is 24 hours and a typical design value is 36 hours. Source: Jack de la Vergne •The main pumps should be placed close to the sump so that the separation will allow for a minimum straight run of pipe equal to five times (preferably ten times) the diameter of the pipe. Source: Various •Turbulence will be sufficient to ensure good mixing of a flocculating agent if the water velocity is at least 1m/s and maintained for 30 seconds in a feed pipe or channel. Source: NMERI of South Africa
    Area: Design •Piping for long runs should be selected on the basis that the water velocity in the pipe will be near 10 feet/sec (3m/s). The speed may be increased up to 50% in short runs. Source: Various •In underground mines, static head is the significant factor for pump design if the pipes are sized properly. To obtain the total head, 5 -10% may be added to the static head to account for all the friction losses without sacrificing accuracy. Source: Andy Pitz •Pump stations for a deep mine served by centrifugal pumps are most economically placed at approximately 2,000-foot (600m) intervals. Source: Andy Pitz •The outlet velocity of a centrifugal pump should be between 10 and 15 feet per second to be economical. Source: Queen's University •Centrifugal pumps should not operate at a speed exceeding 1,800 RPM (except for temporary or small pumps that may operate at 3,600 RPM). This is because impeller wear is proportional to the 2.5 power of the speed. In other words, half the speed means nearly six times the impeller life. Source: Canadian Mine Journal •The maximum lift of a centrifugal pump is a function of the motor torque, which in turn is a function of the supply voltage. Since it is a squared function, a 10% drop in line voltage can result in a 20% loss in head. Source: Jack de la Vergne •The velocity of dirty water being pumped should be greater than 2 fps in vertical piping and 5 fps in horizontal piping. These speeds are recommended to inhibit solids from settling. Source: GEHO •Slime particles less than 5m in diameter cannot be precipitated without use of a flocculating agent. Source: B. N. Soutar

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    Mining Field: Crushers and Rockbreakers
    Area: Crusher Selection •For a hard rock mine application below 600 tonnes/hour, select a jaw as the primary crusher. Over 1,000 tph, select a gyratory crusher. Between these capacities, you have a choice. Source: Chris Ottergren •For a hard rock mine application below 540 tonnes/hour, a jaw crusher is more economical. Above 725 tonnes/hour, jaw crushers cannot compete with gyratory crushers at normal settings (6 -10 inches). Source: Lewis, Cobourn and Bhappu •For an underground hard rock mine, a gyratory crusher may be more economical in the case where its required daily production exceeds 8,000 tonnes of ore. Source: Jack de la Vergne •If the hourly tonnage to be crushed divided by the square of the required gape in inches is less than 0.115, use a jaw crusher; otherwise use a gyratory. (If the required capacity in metric tph is less than 162 times the square of the gape in metres, use a jaw crusher.) Source: Arthur Taggart •Nearly all crushers produce a product that is 40% finer than one-half the crusher setting. Source: Babu and Cook •The product of a jaw crusher will have a size distribution such that the -80% fraction size (d80) is slightly less than the open-side setting of the crusher. For example, if the open-side setting is 6 inches, the d80 product size will be 5¾ inches. Source: Unknown •In a hard rock mine, the product from a jaw crusher will tend to be slabby, while the product from a gyratory crusher may tend to be blocky, the latter being easier to convey through transfer points on a conveyor system. Source: Heinz Schober •Impact crushers (rotary or hammer mills) have the capacity for high reduction ratios (up to 40:1), but are rarely applied to hard rock mines. Since they depend on high velocities for crushing, wear is greater than for jaw or gyratory crushers. Hence, they should not be used in hard rock mines that normally have ores containing more than 15% silica (or any ores that are abrasive). Source: Barry Wills
    Area: Crusher Design •The approximate capacity of a jaw crusher for hard rock application at a typical setting may be obtained by multiplying the width by 10 to get tonnes per hour. For example, a 48 by 60 crusher will have a capacity in the order of 600 tph when crushing ore in a hard rock mine. Source: Jack de la Vergne •The capacity of a jaw crusher selected for underground service should be sufficient to crush the daily requirement in 12 hours. Source: Dejan Polak •For most applications, 7:1 is the maximum practical reduction factor (ratio) for a jaw crusher, but 6:1 represents better design practice. Source: Jack de la Vergne •For most applications, 6:1 is the maximum practical reduction factor (ratio) for a cone crusher, but 5:1 represents better design practice. Source: Jack de la Vergne •Corrugated liner plates designed for jaw crushers (to avoid a slabby product) result in shortening liner life by up to two-thirds and they are more prone to plugging than smooth jaws. Source: Ron Doyle
    Area: Crusher Installation •The crushed ore surge pocket beneath a gyratory crusher should have a live load capacity equal to 20 minutes of crusher capacity or the capacity of two pit trucks. Source: Various •It will take six months to excavate, install, and commission an underground crusher station for a typical jaw crusher. For a very large jaw crusher or a gyratory crusher, it can take nine months. Source: Jim Redpath •The desired grizzly opening for an underground jaw crusher is equal to 80% of the gape of the crusher. Source: Jack de la Vergne •The combination of a jaw crusher and a scalping grizzly will have 15% more capacity than a stand-alone jaw crusher. Source: Ron Casson •As a rule, scalping grizzlies are rarely used anymore for (large) primary crushers. The exception is when ore contains wet fines that can cause acute packing in a gyratory crusher. Source: McQuiston and Shoemaker •The product from a jaw crusher will tend to be less slabby and more even-dimensioned without a scalping grizzly, since slabs do not pass through so readily under this circumstance. Source: A. L. Engels •Removal of the scalping grizzly for a primary jaw crusher can cut the liner life by 50%. It also makes it more difficult to clear a jam when the jaws are filled with fines. Source: Ron Doyle
    Area: Rockbreakers •The capacity of a hydraulic rockbreaker is higher (and the operating cost lower) than a pneumatic rockbreaker. For these reasons, most new installations are hydraulic, despite the higher capital cost. Source: John Kelly •For underground production rates less than 2,000 tpd, it may be economical to size the ore underground with rockbreakers only, otherwise, an underground crusher is usually necessary when skip hoisting is employed. Source: John Gilbert •The operating cost for a stand-alone rockbreaker will be approximately 30% higher than it is for a crusher handling the same daily tonnage. Source: John Gilbert •The capacity of one rockbreaker on a grizzly with the standard opening (± 16 by 18 inches) is in the order of 1,500-2,000 tpd. Source: John Gilbert. •For skips that fit into a standard 6 by 6 shaft compartment, the maximum particle size that is normally desired for skip hoisting is obtained when run-of-mine muck has been passed through a grizzly with a 16-18 inch opening. Skips hoisted in narrow shaft compartments may require a 12-14 inch spacing, while oversize skips may handle muck that has passed a 24-30 inch spacing. Source: Jack de la Vergne •A pedestal mounted rockbreaker installed should be equipped with a boom that enables a reach of 20 feet (6m). Source: Peter van Schaayk

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    جديد


    تاريخ التسجيل: Oct 2011
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    alshangiti
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    Thumbs Up
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    Given: 8
    Mining Field: Ventilation and Air Conditioning
    Area: General •An underground trackless mine may require 10 tons of fresh air to be circulated for each ton of ore extracted. The hottest and deepest mines may use up to 20 tons of air for each ton of ore mined. Source: Northern Miner Press •A factor of 100 cfm per ore-ton mined per day can be used to determine preliminary ventilation quantity requirements for most underground mining methods. Hot mines using ventilation air for cooling and mines with heavy diesel equipment usage require more air. Uranium mines require significantly higher ventilation quantities, up to 500 cfm per ton per day. Block cave and large-scale room and pillar mining operations require significantly lower ventilation quantities, in the range of 20 to 40 cfm per ton per day for preliminary calculations. Source: Scott McIntosh •Ventilation is typically responsible for 40% of an underground mine's electrical power consumption. Source: CANMET •If the exhaust airway is remote from the fresh air entry, approximately 85% of the fresh air will reach the intended destinations. If the exhaust airway is near to the fresh air entry, this can be reduced to 75%, or less. The losses are mainly due to leaks in ducts, bulkheads, and ventilation doors. Source: Jack de la Vergne •Mine Resistance - For purposes of preliminary calculations, the resistance across the mine workings between main airway terminals underground (shafts, raises, air drifts, etc.) may be taken equal to one-inch water gauge. Source: Richard Masuda •Natural pressure may be estimated at 0.03 inches of water gage per 10 degrees Fahrenheit difference per 100 feet difference in elevation (at standard air density). Source: Robert Peele
    Area: Airways •The maximum practical velocity for ventilation air in a circular concrete production shaft equipped with fixed (rigid) guides is 2,500 fpm (12.7m/s). Source: Richard Masuda •The economic velocity for ventilation air in a circular concrete production shaft equipped with fixed (rigid) guides is 2,400 fpm (12m/s). If the shaft incorporates a man-way compartment (ladder way) the economic velocity is reduced to about 1,400 fpm (7m/s). Source: A.W.T. Barenbrug •The maximum velocity that should be contemplated for ventilation air in a circular concrete production shaft equipped with rope guides is 2,000 fpm and the recommended maximum relative velocity between skips and airflow is 6,000 fpm. Source: Malcom McPherson •The "not-to-exceed" velocity for ventilation air in a bald circular concrete ventilation shaft is 4,000 fpm (20m/s). Source: Malcom McPherson •The typical velocity for ventilation air in a bald circular concrete ventilation shaft or a bored raise is in the order of 3,200 fpm (16m/s) to be economical and the friction factor, k, is normally between 20 and 25. Source: Jack de la Vergne •The typical velocity for ventilation air in a large raw (unlined) ventilation raise or shaft is in the order of 2,200 fpm (11m/s) to be economical and the friction factor, k, is typically between 60 and 75. Source: Jack de la Vergne •The typical range of ventilation air velocities found in a conveyor decline or drift is between 500 and 1,000 fpm. It is higher if the flow is in the direction of conveyor travel and is lower against it. Source: Floyd Bossard •The maximum velocity at draw points and dumps is 1,200 fpm (6m/s) to avoid dust entrainment. Source: John Shilabeer •A protuberance into a smooth airway will typically provide four to five times the resistance to airflow as will an indent of the same dimensions. Source: van den Bosch and Drummond •The friction factor, k, is theoretically constant for the same roughness of wall in an airway, regardless of its size. In fact, the factor is slightly decreased when the cross-section is large. Source: George Stewart
    Area: Ducts •For bag duct, limiting static pressure to approximately 8 inches water gage will restrict leakage to a reasonable level. Source: Bart Gilbert •The head loss of ventilation air flowing around a corner in a duct is reduced to 10% of the velocity head with good design. For bends up to 30 degrees, a standard circular arc elbow is satisfactory. For bends over 30 degrees, the radius of curvature of the elbow should be three times the diameter of the duct unless turning vanes inside the duct are employed. Source: H.S. Fowler •The flow of ventilation air in a duct that is contracted will remain stable because the air-flow velocity is accelerating. The flow of ventilation air in a duct that is enlarged in size will be unstable unless the expansion is abrupt (high head loss) or it is coned at an angle of not more than 10 degrees (low head loss). Source: H. S. Fowler
    Area: Fans •Increasing fan speed by 10% may increase the quantity of air by 10%, but the power requirement will increase by 33%. Source: Chris Hall •The proper design of an evasée (fan outlet) requires that the angle of divergence not exceed 7 degrees. Source: William Kennedy
    Area: Air Surveys •For a barometric survey, the correction factor for altitude may be assumed to be 1.11 kPa/100m (13.6 inches water gage per thousand feet). Source: J.H. Quilliam
    Area: Clearing Smoke •The fumes from blasting operations cannot be removed from a stope or heading at a ventilation velocity less than 25 fpm (0.13m/s). A 30% higher air velocity is normally required to clear a stope. At least a 100% higher velocity is required to efficiently clear a long heading. Source: William Meakin •The outlet of a ventilation duct in a development heading should be advanced to within 20 duct diameters of the face to ensure it is properly swept with fresh air. Source: J.P. Vergunst •For sinking shallow shafts, the minimum return air velocity to clear smoke in a reasonable period of time is 50 fpm (0.25m/s). Source: Richard Masuda •For sinking deep shafts, the minimum return air velocity to clear smoke in a reasonable period of time is 100 fpm (0.50m/s). Source: Jack de la Vergne •For sinking very deep shafts, it is usually not practical to wait for smoke to clear. Normally, the first bucket of men returning to the bottom is lowered (rapidly) through the smoke. Source: Morris Medd
    Area: Mine Air Heating •To avoid icing during winter months, a downcast hoisting shaft should have the air heated to at least 50C. (410 F.). A fresh air raise needs only 1.50C. (350 F.). Source: Julian Kresowaty •When calculating the efficiency of heat transfer in a mine air heater, the following efficiencies may be assumed.
    90% for a direct fired heater using propane, natural gas or electricity
    80% for indirect heat transfer using fuel oil
    Source: Various •When the mine air is heated directly, it is important to maintain a minimum air stream velocity of approximately 2,400 fpm across the burners for efficient heat transfer. If the burners are equipped with combustion fans, lower air speeds (1,000 fpm) can be used. Source: Andy Pitz •When the mine air is heated electrically, it is important to maintain a minimum air stream velocity of 400 fpm across the heaters. Otherwise, the elements will overheat and can burn out. Source: Ed Summers
    Area: Heat Load •The lowest accident rates have been related to men working at temperatures below 70 degrees F and the highest to temperatures of 80 degrees and over. Source: MSHA •Auto compression raises the dry bulb temperature of air by about 1 degree Celsius for every 100m the air travels down a dry shaft. (Less in a wet shaft.) The wet bulb temperature rises by approximately half this amount. Source: Various •At depths greater than 2,000m, the heat load (due to auto compression) in the incoming air presents a severe problem. At these depths, refrigeration is required to remove the heat load in the fresh air as well as to remove the geothermal heat pick-up. Source: Noel Joughin •At a rock temperature of 50 degrees Celsius, the heat load into a room and pillar stope is about 2.5 kW per square meter of face. Source: Noel Joughin •In a hot mine, the heat generated by the wall rocks of permanent airways decays exponentially with time – after several months it is nearly zero. There remains some heat generated in permanent horizontal airways due to friction between the air and the walls. Source: Jack de la Vergne •A diesel engine produces 200 cubic feet of exhaust gases per Lb. of fuel burned and consumption is approximately 0.45 Lb. of fuel per horsepower-hour. Source: Caterpillar and others •Normally, the diesel engine on an LHD unit does not run at full load capacity (horsepower rating); it is more in the region of 50%, on average. In practice, all the power produced by the diesel engines of a mobile equipment fleet is converted into heat and each horsepower utilized produces heat equivalent to 42.4 BTU per minute. Source: A.W.T.Barenbrug •The heat load from an underground truck or LHD is approximately 2.6 times as much for a diesel engine drive as it is for electric. Source: John Marks •The efficiency of a diesel engine can be as high as 40% at rated RPM and full load, while that of an electric motor to replace it is as high as 96% at full load capacity. In both cases, the efficiency is reduced when operating at less than full load. Source: Various •Normally, the electric motor on an underground ventilation fan is sized to run at near full load capacity and it is running 100% of the time. In practice, all the power produced by the electric motor of a booster fan or development heading fan is converted into heat and each horsepower (33,000 foot-Lb./minute) produces heat equivalent to 42.4 BTU per minute. (1 BTU = 778 foot-Lbs.) Source: Jack de la Vergne •Normally, the electric motor on a surface ventilation fan is sized to run at near full load capacity and it is running 100% of the time. In practice, about 60% of the power produced by the electric motors of all the surface ventilation fans (intake and exhaust) is used to overcome friction in the intake airways and mine workings (final exhaust airways are not considered). Each horsepower lost to friction (i.e. static head) is converted into heat underground. Source: Jack de la Vergne •Heat generated by electrically powered machinery underground is equal to the total power minus the motive power absorbed in useful work. The only energy consumed by electric motors that does not result in heat is that expended in work against gravity, such as hoisting, conveying up grade, or pumping to a higher elevation. Source: Laird and Harris
    Area: Air Conditioning and Refrigeration •In the Republic of South Africa, cooling is required when the natural rock temperature reaches the temperature of the human body (98.6 degrees F). Source: A.W.T. Barenbrug •A rough approximation of the cooling capacity required for a hot mine in North America is that the TR required per ton mined per day is 0.025 times the difference between the natural rock temperature (VRT) and 95 degrees F. For example, a 2,000 ton per day mine with a VRT of 140 degrees F. at the mean mining depth will require approximately 0.025 x 45 x 2,000 = 2,250 TR. Source: Jack de la Vergne •The cold well (surge tank) for chilled surface water should have a capacity equal to the consumption of one shift underground. Source: J. van der Walt •At the Homestake mine, the cost of mechanical refrigeration was approximately equal to the cost of ventilation. Source: John Marks

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  7. [17]
    alshangiti
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    تاريخ التسجيل: Mar 2007
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    Thumbs Up
    Received: 89
    Given: 8
    Mineral Economics
    Area: Metal Price•The long-term average price of a common mineral commodity (the price best used for economic evaluation in a feasibility study) is 1.5 times the average cost of production, worldwide. Source: Sir Ronald Prain
    Area: Pre-production Capital Cost•The pre-production capital cost estimate (Capex) should include all construction and operating expenses until the mine has reached full production capacity or three months after reaching 50% of full capacity, whichever occurs first. This is the basic transition point between capital and operating costs. Source: John Halls•The pre-production capital cost expenditure includes all costs of construction and mine development until three months after the mine has reached 25% of its rated production capacity. Source: Jon Gill
    Area: Cash Flow•The total cash flow must be sufficient to repay the capital cost at least twice. Source: L. D. Smith•Project loans should be repaid before half the known reserves are consumed. Source: G.R Castle•Incremented cash flow projections should each be at least 150% of the loan repayment scheduled for the same period. Source: G.R. Castle•The operating cost should not exceed half the market value of minerals recovered. Source: Alan Provost
    Area: Net Present Value•The discount factor employed to determine the NPV is often 10%; however, it should be Prime + 5%. Source: G.R. Castle•In numerous conversations with managers of mining firms, I have found that 15% in real terms is the common discount rate used for decision purposes. Source: Herbert Drecshler (1980)•The true present value (market value) of a project determined for purposes of joint venture or outright purchase is equal to half the NPV typically calculated. Source: J. B. Redpath
    Area: Rate of Return•The feasibility study for a hard rock mine should demonstrate an internal rate of return (IRR) of at least 20% – more during periods of high inflation. Source: J. B. Redpath
    Area: Working Capital•Working capital equals ten weeks operating cost plus cost of capital spares and parts. Source: Alan O'Hara
    Area: Closure Costs•The salvage value of plant and equipment should pay for the mine closure costs. Source: Ron Haflidson

    0 Not allowed!


    المهندس / يحيى بن محمد الشنقيطى

  8. [18]
    alshangiti
    alshangiti غير متواجد حالياً

    مشرف وإستشاري هندسة المناجم


    الصورة الرمزية alshangiti


    تاريخ التسجيل: Mar 2007
    المشاركات: 1,472

    وسام مشرف متميز

    Thumbs Up
    Received: 89
    Given: 8
    * • An allowance (such as 15%) should be specifically determined and added to the contractor's formal bid price for a mining project to account for contract clauses relating to unavoidable extra work, delays, ground conditions, over-break, grouting, de-watering, claims, and other unforeseen items. Source: Jack de la Vergne
    *
    Area: Engineering, Procurement, and Construction Management
    * • The Engineering, Procurement, and Construction Management (EPCM) cost will be approximately 17% for surface and underground construction and 5% for underground development. Source: Jack de la Vergne
    *
    Area: Overbreak
    * • The amount of over-break to be estimated against rock for a concrete pour will average approximately one foot in every applicable direction, more at brows, lips, and in bad ground. Source: Jack de la Vergne
    * • On average, for each one cubic yard of concrete measured from the neat lines on drawings, there will be two cubic yards required underground, due to over-break and waste. Source: Jack de la Vergne
    *
    Area: Haulage
    * • The economical tramming distance for a 5 cubic yard capacity LHD is 500 feet and will produce 500 tons per shift, for an 8-yard LHD, it is 800 feet and 800 tons per shift. Source: Sandy Watson
    * • Haulage costs for open pit are at least 40% of the total mining costs; therefore, proximity of the waste dumps to the rim of the pit is of great importance. Source: Frank Kaeschager
    *
    Area: Miscellaneous
    * • The installed cost of a long conveyorway is approximately equal to the cost of driving the drift or decline in which it is to be placed. Source: Jack de la Vergne
    * • In a trackless mine operating around the clock, there should be 0.8 journeyman mechanic or electrician on the payroll for each major unit of mobile equipment in the underground fleet. Source: John Gilbert
    * • On average, for each cubic yard of concrete measured from the neat lines on drawings, approximately 110 Lbs. of reinforcing steel and 12 square feet of forms will be required. Source: Jack de la Vergne
    * • The overall advance rate of a trackless heading may be increased by 30% and the unit cost decreased by 15% when two headings become available. Source: Bruce Lang
    * • The cost to slash a trackless heading wider while it is being advanced is 80% of the cost of the heading itself, on a volumetric basis. Source: Bruce Lang

    0 Not allowed!


    المهندس / يحيى بن محمد الشنقيطى

  9. [19]
    alshangiti
    alshangiti غير متواجد حالياً

    مشرف وإستشاري هندسة المناجم


    الصورة الرمزية alshangiti


    تاريخ التسجيل: Mar 2007
    المشاركات: 1,472

    وسام مشرف متميز

    Thumbs Up
    Received: 89
    Given: 8
    Mining Field: Cost Estimating
    *
    Area: Cost of Estimating
    * • A detailed estimate for routine, repetitive work (i.e. a long drive on a mine level) may cost as little as 0.5% of the project cost. On the other hand, it may cost up to 5% to adequately estimate projects involving specialized work, such as underground construction and equipment installation. Source: Various
    *
    Area: Cost of Feasibility Study
    * • The cost of a detailed feasibility study will be in a range from 0.5% to 1.5% of the total estimated project cost. Source: Frohling and Lewis
    * • The cost of a detailed or "bankable" feasibility study is typically in the range of 2% to 5% of the project, if the costs of additional (in-fill) drilling, assaying, metallurgical testing, geotechnical investigations, etc. are added to the direct and indirect costs of the study itself. Source: R. S. Frew
    *
    Area: Budget Estimates
    * • An allowance (such as 15%) should be specifically determined and added to the contractor's formal bid price for a mining project to account for contract clauses relating to unavoidable extra work, delays, ground conditions, over-break, grouting, de-watering, claims, and other unforeseen items. Source: Jack de la Vergne

    0 Not allowed!


    المهندس / يحيى بن محمد الشنقيطى

  10. [20]
    alshangiti
    alshangiti غير متواجد حالياً

    مشرف وإستشاري هندسة المناجم


    الصورة الرمزية alshangiti


    تاريخ التسجيل: Mar 2007
    المشاركات: 1,472

    وسام مشرف متميز

    Thumbs Up
    Received: 89
    Given: 8
    Back fill.


    Mining Field: Backfill
    *
    Area: General
    * • The cost of backfilling will be near 20% of the total underground operating cost. Source: Bob Rappolt
    * • The capital cost of a paste fill plant installation is approximately twice the cost of a conventional hydraulic fill plant of the same capacity. Source: Barrett, Fuller, and Miller
    * • If a mine backfills all production stopes to avoid significant delays in ore production, the daily capacity of the backfill system should be should be at least 1.25 times the average daily mining rate (expressed in terms of volume). Source: Robert Currie
    * • The typical requirement for backfill is approximately 50% of the tonnage mined. It is theoretically about 60%, but all stopes are not completely filled and tertiary stopes may not be filled at all. Source: Ross Gowan
    * • It is common to measure the strength of cemented backfill as if it were concrete (i.e. 28 days), probably because this time coincides with the planned stope turn-around cycle. Here it should be noted that while concrete obtains over 80% of its long- term strength at 28 days, cemented fill might only obtain 50%. In other words, a structural fill may have almost twice the strength at 90 days as it had at 28 days. Source: Jack de la Vergne
    *
    Area: Hydraulic Fill
    * • Because the density of hydraulic fill when placed is only about half that of ore, unless half the tailings can be recovered to meet gradation requirements, a supplementary or substitute source of fill material is required. Source: E. G. Thomas
    *
    Area: Cemented Rock Fill
    * • A 6% binder will give almost the same CRF strength in 14 days that a 5% binder will give in 28 days. This rule is useful to know when a faster stope turn-around time becomes necessary. Source: Joel Rheault
    * • As the fly ash content of a CRF slurry is increased above 50%, the strength of the backfill drops rapidly and the curing time increases dramatically. A binder consisting of 35% fly ash and 65% cement is deemed to be the optimal mix. Source: Joel Rheault
    * • The size of water flush for a CRF slurry line should be 4,000 US gallons. Source: George Greer
    * • The optimum W/C ratio for a CRF slurry is 0.8:1, but in practice, the water content may have to be reduced when the rock is wet due to ice and snow content of quarried rock or ground water seepage into the fill raise. Source: Finland Tech
    * • The actual strength of CRF placed in a mine will be approximately 2/3 the laboratory value that is obtained from standard 6 inch diameter concrete test cylinders, but will be about 90% of the value obtained from 12 inch diameter cylinders. Source: Thiann Yu
    *
    Area: Paste Fill
    * • Only about 60% of mill tailings can be used for paste fill over the life of a mine because of the volume increase, which occurs as a result of breaking and comminuting the ore. Source: David Landriault
    * • Experience to date at the Golden Giant mine indicates that only 46% of the tailings produced can be used for paste fill. Source: Jim Paynter
    * • Very precise control of pulp density is required for gravity flow of paste fill. A small (1-2%) increase in pulp density can more than double pipeline pressures (and resistance to flow). Source: David Landriault

    0 Not allowed!


    المهندس / يحيى بن محمد الشنقيطى

  
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