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By J. H. Elledge

Hydraulic Transportation of Phosphate Ore from, a 40,000 Cubic Yard per Day Strip Mine IMC is transporting 1500 yards per hour of phosphate ore in a water slurry a distance of approximately five miles from a strip mine to a flotation beneficiation plant at their Noralyn Operation. in Barrow, Florida. This ore, which consists of approximately 1/4 phosphate, 1/2 silica sand, and 1/4 clay slimes, is being pumped through a 20 inch O. D. pipeline using seven 20 and 18 inch GIW metal pumps direct connected to electric, motors. This paper details some of the problems involved and the specific methods used to accomplish this bulk material handling requirement.

Jan 1, 1966

By G. D. MacDonald

The Grasberg porphyry copper-gold deposit of P.T. Freeport Indonesia, Co. is located in the Jayawijaya Mountains of Irian Jaya, Indonesia approximately six km west northwest of Puncak Jaya, the highest point on the island of New Guinea. Disseminated mineralization is confined to the Grasberg intrusive complex, a multiphase intrusion of dioritic to monzonitic composition which intruded tightly folded carbonate rocks of the Tertiary Faumai and Ainod formations. The high tonnage and grade of the deposit (1 X 109 tonnes at 1.4 % copper and 1. 8 ppm gold) appear to result from a succession of coincident intrusive and mineralizing events about three million years ago. This paper will focus on the history and character of these events.

Jan 1, 1993

By B. Yarar

The critical surface tension of wetting (?c) of Cu2S, FeS2, PbS, and MoS2 were determined by plotting flotation recovery versus YLV which was arranged by methanol, and Extra-polating to % R = 0. ?c values ranged from 26 dyne cm-1 for MoS2 to 49 dyne cm-1 for PbS and elemental sulfur gave a value of 31.5 dyne cm-1. Minerals exhibiting natural flotability of different extents, could be separated from their mixtures in the absence of conventional flotation reagents, but controlled ?LV. Correlations between non-stoichiometry and degree of natural hydrophobicity for MOS2 are also reported in this paper.

Jan 1, 1983

By J. H. White

It is hypothesized that the Hotelling Valuation Principle (HVP) can determine the value of developed gold reserves using simple methods and easily obtained data. This paper reports the results of an empirical test of this hypothesis. Using publicly available data from North American gold companies, we obtained data for V, R, p, and c and ran a simple linear regression using equation (1.1) as the definition of the dependent and explanatory variables. By examining the value of the coefficients of the resulting regression equation, we can draw conclusion as to whether or not the HVP can be confident1y used as an estimator of the value of developed gold reserves.

Jan 1, 1995

By Shawn Benner, David Blowes, Carol Ptacek

A full-scale reactive barrier was installed into an aquifer impacted by acid mine drainage in August 1995 at the Nickel Rim mine site near Sudbury, Ontario. The primary mechanism of water quality improvement is bacterially mediated sulfate reduction and metal sulfide precipitation. Dramatic changes in the groundwater chemistry are observed with up to 90% decline in Fe (from >18 to <2 mmol L-1) and increases in alkalinity from <1 to >10 mmol L-1 (as CaCO3). Over a 3 year period, the rate of Fe removal declines by approximately 35%. Primary controls on barrier performance are organic carbon reactivity, groundwater temperature, and variable residence time within the barrier.

Jan 1, 2000

By J. C. Franklin

The U.S. Department of the Interior, Bureau of Mines, has developed a data acquisition system to continuously monitor radon, working level, absolute pressure, temperature, relative humidity, and air velocity. Data have been taken in intake and exhaust drifts, working stopes, and sealed-off areas of positive and negative ventilated uranium mines. Radon emanation rates, ranging from 1.7- to 12.2 x 2 10-l4 Ci/cm /sec were determined in four mines. The ratios of short duration, primary fan shutdown time to radon normalization time were 1 to 2 and 1 to 4, respectively, for-negatively and positively ventilated mines. Blasting and slushing caused radon increases of 160 and 67 percent, respectively.

Jan 1, 1980

By E. D. Gardner, William O. Vandenburg

This chapter on the square-set system of mining em¬bodies the results of a special study of the subject made as a part of the US Bureau of Mines (USBM) general investigation of mining methods and costs. Data for the chapter were gathered by the authors during visits to mines using square-set mining, from information circu¬lars prepared by consulting engineers to the USBM on mining methods at individual mines, and from articles in the technical press and the Transactions of the American Institute of Mining and Metallurgical Engineers. Al¬though some of the properties discussed are no longer in operation, the methods described are still valid. The square-set system of mining has been very widely applied and is one of the oldest methods in use in the US. It is mainly employed for extracting rela¬tively high-grade ore in heavy ground. It is also useful in taking out pillars left in mining by other methods. It is used to some extent as an auxiliary method with many underground systems and as an alternate with cut-and¬fill, shrinkage, and open-stope methods when heavy ground is encountered. In some cases as much as 50% of the ore may be extracted by square setting at mines listed as using some other method of mining. A large tonnage has been mined yearly by square setting, prin¬cipally in the western states. Due to the high cost of timber and the relatively low average production per manshift, this system generally is not used when the ore can be mined by some other method; it usually is adopted as a last resort. Square setting, however, has a special field in that ore bodies not minable by other methods can be stoped by this system. As square-set mining was widely known, there was a tendency for years to use the method regardless of its suitability. With the increasing need for economy and the development of other methods, square setting has been superseded in many mines by cut-and-fill, top slic¬ing, or sublevel caving methods. The term square-set mining as used in this chapter includes all stoping methods in which the space made by the removal of the ore is timbered with framed sets. The sets usually are square in horizontal section. The stope may or may not be filled with waste rock as stoping progresses. HISTORY OF THE SQUARE-SET SYSTEM Prior to 1860, open-stope methods with stulls for casual support accounted for most of the mineral pro¬duction from underground mines. Attempts to mine large ore bodies that required artificial support resulted in disastrous cave-ins and loss of life. The art of mining received a tremendous impetus following the introduc¬tion of methods for mining the bonanza ore body which was discovered in 1860 in the Ophir mine on the Comstock lode, Virginia City, NV. This ore body was phe¬nomenally rich and eventually proved to be 122 m (400 ft) high, 27 m (90 ft) wide, and 98 m (320 ft) long. The mining of this large high-grade ore body taxed the best engineering talent of the time; the problem was finally solved by Philip Deidesheimer. According to Eliot Lord (1882): "At the 50-foot level the vein of black sulphurets was only 3 or 4 feet thick, and could be readily extracted through a drift along its line, propping up the walls and roof when necessary by simple uprights and cap-posts. As the ledge descended the sulphuret vein grew broader until, at the depth of 175 feet, it was 65 feet in width, and the miners were at a loss how to proceed, for the ore was so soft and crumbling that pillars could not be used to support the roof as in coal mines. They spliced timbers together to hold up the caving ground; but these jointed props were too weak and ill-supported to with¬stand the pressure upon them, and were constantly broken and thrown out of place. The dilemma was a curious one. Surrounded by riches, they were yet un¬able to carry them off and their mass of black sulphurets bade fair to become a white elephant on their hands. The Ophir Company began to wish themselves less for¬tunate, as their miners narrowly escaped burial day after day in their attempts to stope out the ore. "A young mining engineer, Philip Deidesheimer, was in charge of a quartz mine in Georgetown, El Dorado County, Calif., in the autumn of 1860, when this serious check to the development of the lode occurred. At the request of William F. Babcock, a trustee of the Ophir Company, Mr. Deidesheimer left his California mine and crossed the mountains with a letter from the direc¬tors of the Ophir Company, authorizing him to inspect the workings of their mines and make such changes in the method of timbering as should seem to him expe¬dient. After examining the vein he designed, in the course of a few weeks, a system of timbering which proved to be exactly adapted to the requirements of the work. Experiments which he had previously made in California gravel and quartz mines furnished the out¬line of his plan. This was to frame timbers together in rectangular sets, each set being composed of a square base, placed horizontally, formed of four timbers, sills, and crosspieces from 4 to 6 feet long, surmounted at the corners by four posts from 6 to 7 feet high, and capped by a framework similar to the base. The cap pieces forming the top of any set were at the same time the sills or base of the next set above. These sets could readily be extended to any required height and over any given area, forming a series of horizontal floors, built up from the bottom sets like the successive stories of a house. The spaces between the timbers were filled with waste rock or with wooden braces, forming a solid cube when¬ever the maximum degree of firmness was desired. By adjustments of these sets the ore bodies along the line of the lode were extracted with comparative ease and security." Most mining methods are evolved slowly and are the result of the combined efforts of a number of minds; Deidesheimer&apos;s feat was remarkable in that his system was immediately successful. He died in 1916 at the age

Jan 1, 1982

By John W. Amos, John A. Christophersen

The Deister No. 7, "77" and "88" Coal Washing Tables have been used by the coal industry in the cleaning of finer sizes of coal (-1/2") since washing of coals became necessary. This paper will deal with the in-plant use of the CONCENCO "88" Deister Coal Washing Table, along with a brief description of the Deister Refuse Superscalper. Efficient in-plant operation of the Deister Coal Washing Table can be accomplished by the adjustment of six basic variables. These variables are: 1. Headmotion 2. Feed 3. Riffle Pattern 4. Dressing Water 5. End Elevation Adjustment 6. Side Tilt Adjustment VARIABLE NO. 1 HEADMOTION The reciprocating movement of the "88" Table is developed within the headmotion entirely by rotating eccentric weights. Four parallel shafts each carry pairs of eccentric weights (Figure 1-A). The shafts are geared together with gears of 2 to 1 and 1 to 1 ratios, so that the two shafts in vertical alignment which carry the small eccentric weights operate at twice the speed of the other two shafts. The other two shafts are also in vertical alignment with each other and carry large eccentric weights. As power is supplied to one high-speed shaft, once each revolution the forces produced by the large eccentric weights are reinforced and then opposed by the small eccentric weights in a horizontal plane. The direction of movement of the table decks is reversed much more sharply at the forward end than at the backboard end of the stroke, causing material to move toward the refuse end of the table. Because in the vertical plane all centrifugal forces oppose each other and cancel out, there is no vertical movement and virtually no impact to the support steel. The "88" Coal Washing Table is suspended by wire-rope cables from three overhead steel beams, and operates with a perfectly uniform and exact differential motion curve in every stroke (Figure 1-B). The decks, being attached to the headmotion, are reciprocated in a parallel line from the head- motion to the refuse corner at speeds varying from 260-300 strokes per minute. Both length of stroke and speed are set at the factory for the coal characteristic desired and should not be changed. A 3 HP motor is mounted on the top of the headmotion and uses 3 V-belts to transmit driving power to the headmotion. The headmotion must be kept running on a horizontal plane. Punch marks are provided on the side and end of the main frame for use in leveling the motion (Figure 1-C). In leveling the motion, the work must always be done with two cables. To level the long side of the headmotion both front or both rear cables are raised or lowered to bring motion back to level. To level side to side, both right side cables or left side cables are raised or lowered to bring motion back to level VARlABLE NO. 2 FEED When tabling different size coals, each deck of a "88" DIAGONAL-DECK Coal Washing Table will handle feeds as described in figure 2-A.

Jan 1, 1988

By Brian G. White

The Coeur d&apos;Alene Mining District of northern Idaho is remarkable for its number of large ore bodies and for the amount of its historical production of lead, zinc and silver (Fig. 1). Since veins were first discovered in 1884, more than 127 Mt (140 million st) of ore have been produced from narrow, high-grade veins. Mining continues at the Lucky Friday, Galena and Sunshine mines (Fig. 2). All have been in production for decades. At these mines, more efficient mining and exploration techniques have reduced costs while adding significant new reserves. Even the old Bunker Hill Mine, where ore was first found in 1875, continues operation on a limited basis, partly involving production of world-class pyromorphite specimens. Until recently, the Coeur Mine was also a profitable silver producer. Other large producers have included the Star-Morning, Hecla, Page, Standard-Mammoth, Hercules and Gold Hunter mines (Fig. 2). The first deposits were discovered by prospectors who spilled over from the Murray gold mining district, 24 km (15 miles) to the north. The Bunker Hill Mine was founded the year after the first veins were discovered. Bunker Hill has now operated for more than 100 years. Production includes 2.3 Mt (2.6 million st) of lead, 1.1 Mt (1.2 million st) of zinc and 4.1 kt (132 million oz) of silver. Total district production through 1996 (Springer, 1997) amounted to 7.3 Mt (8.1 million st) of lead, 3 Mt (3.3 million st) of zinc, and 34.5 kt (1.109 billion oz) of silver. This record establishes the district as the largest recorded producer of silver in the world, while also yielding major quantities of lead and zinc. Only Potosi, Bolivia, may have produced more silver, although records are unclear. Despite the district&apos;s longevity, historic production and current activity, the published literature on the district lags behind with regard to recent concepts of district geology. The most recent comprehensive US Geological Sur- vey (USGS) bulletins (Hobbs et al., 1965; Fryklund, 1964) are dated. Much research in the last half century has been short-lived or geographically restricted because the district is large and its geology is complex - and even this research tends to remain unpublished. Thus, it has been difficult for in- [ ]

Jan 1, 1998

By J. C. Stankus, R. Ochsner, B. Hanson

"Bull Mountains No. 1 mine of Signal Peak Energy made another success in its 4 Right pre-developed recovery entry on March 23 and 24, 2015, after the first successful cut-through and shield retrieval of 3 Right pre-developed longwall recovery entry in 2014[1]. Under the same geological conditions, the major difference of 4 Right and 3 Right is the 600 psi Unconfined Compressive Strength (UCS) value of light-weight concrete in 4 Right versus 800 psi in 3 Right. In addition, one more recovery chute was added in 4 Right. The unique geological conditions in this mine include: (1) the low overburden of approximately 200-260 ft, (2) prevalence of roof joints parallel to the face in the sandstone/siltstone dominated overburden, (3) proximity of the Rehder rider seam to the bench Mammoth coal seam, and (4) prompt surface subsidence with the face advancement. The 4 Right recovery entry was 42 ft wide and developed in two stages, 20 ft wide in the inby Phase I, and 22 ft wide in the outby Phase II. The entry roof was reinforced with steel wire mesh/non-metallic recovery mesh, primary bolts, and high-capacity supplemental SUMO cables. After all the support was installed, the entry was completely backfilled with specially designed, cuttable, low-density cement. It took approximately half time to easily cut through the concrete with less dust in 4 Right as compared to the concrete-cut-through time in the 3 Right. The updated study further explored the nature and mechanism of effective roof control approaches in the pre-developed recovery entry of this unique geology mine, aiming at providing the safest working environment for protecting the miners. GEOLOGICAL AND MINING CONDITIONS OF THE RECOVERY ENTRIES Bull Mountains No. 1 mine is located 30 miles north of Billings, Montana, and currently mines bituminous coal from the Mammoth coal seam. The mine employed a conventional recovery face in Panel #1 and pre-developed recovery entries in Panels #2, #3, and #4 (Figure 1). The typical mining and geological conditions are as follows:Overburden Depth The overburden above the Mammoth coal seam in the #1-#4 Right recovery areas is approximately 200-260 ft. This low cover, in combination with specific roof geological features (especially the prevalent roof joints and proximity of rider seam), constitute unique ground movement characteristics when the longwall face advances into the recovery entry if no sufficient roof support is installed."

Jan 1, 2016

By Alden B. Carpenter

Brines (a solution containing more than 100,000 mg/l dissolved solids) in sedimentary basins are genetically related to evaporites in one of three ways: (1) dissolution of evaporite minerals, generally halite; (2) interstitial fluids in evaporites which are expelled from their source rock during compaction; (3) incongruent alteration of hydrous evaporite minerals (carnallite). Most brines of commercial interest originate as interstitial fluids in halite-rich rocks. These fluids are initially rich in magnesium, sulfate, and potassium but evolve in composition to an Na-Ca-Cl brine during their migration to their present host rock. A particularly important reaction in this process is: 2CaCO3 + Mgt+ = CaMg(CO3)2 + Cat (1a) This reaction reduces the magnesium content of the brine and ultimately converts it to a calcium-rich brine. The loss of sulfate occurs by two mechanisms. One of these is by direct precipitation as calcium sulfate (either as gypsum or anhydrite): Ca 2+ + SO4 2- = CaSO4 (2a) Additional sulfate may be lost from the brine by sulfate reduction: S042 + 1.33CH2 + 214- = H2S + 1.33CO2 + 1.33H20 (3a) The loss of carbonate species occurs by precipitation as carbonates: H2O + C02 + Cat = CaCO3 + 2H+ (4a) The net result of reactions 3 and 4 is: Ca 2+ + SO4 2- + 1.33CH2 = H2S + CaCO3 + 0.33H2O + 0.33CO2 (5a) The loss of potassium from brines is caused by the formation of potassium aluminosilicates: 2K+ + Al2Si2O5 (OH)4 + 4SiO2 = 2H + 2KAlSi2O8 + H2O (6a) The hydrogen ions produced by this reaction dissolve carbonates: 2H+ + CaCO3 = Ca 2+ + CO2 + H2O (7a) The net result of reactions 6 and 7 is: 2K+ + CaCO3 + Al2Si205 OH)4 + 4SiO2 = Ca2 + 2KA1Si3O8 + H2O + C02 (8a) A study of brines from central Mississippi has shown that the composition of these brines is related to the interaction of pore fluids from evaporites with minerals in the enclosing sedimentary rocks.

Jan 1, 1978

By John Turner, John A. Gibson, David Cass

Contents A. Introduction B. Description of Robbins Company involvement in several of the English Channel Tunnel projects C. Identification of project requirements and comparison of TBM features D. Description of the manufacturers separate roles in the -235 (T-1) TBM E. T-1 performance to January 15, 1989 F. Summary

Jan 1, 1989

By K. C. Ko, M. D. Bunnell

The studied coal mi1e is an underground coal mine located in the northern Wasatch Plateau of central Utah (Figure 1) . Since operation began in 1981, in situ stress conditions have been encountered which differ from the assumed conditions utilized during early mine design efforts. Geologic structures also varied somewhat from those projected during initial geologic interpretations of the area. The encountered geologic conditions combined with the unusual stress conditions prompted a combined study of in situ stress and stress-related geologic structures. The study was conducted in two phases, with phase 1 consisting of detailed underground mapping and analysis of stress-related geologic structures to determine stress orientations from which they originated; and phase 2, consisting of over- coring two sites within the mine to determine present in situ stress orientations and magnitudes. GEOLOGIC ANALYSIS Three types of geologic structures have been encountered in the mine that act as indicators of stress orientations of the geologic past. Two of these structures, faulting, and igneous dikes, are directly measurable in the mine. A third type of structure, photolineation, was detected through remote sensing methods and shoe god correlation to roof conditions in certain areas of the mine. Faults Two major orientations of faulting have been encountered in the studied coal mine, including a system of N 80 W striking normal faults and a system of N 55 to 70 W striking oblique-slip faults. Only data collected from the oblique-slip faults were used for this study due to the fact that these faults offset the N 80 W faults wherever the two systems intersect, indicating that the N 55 to 70 W system was created by a more recent stress field. Figure 2 shows locations within the mine where data were collected from seven locations along oblique-slip faults. Data collection consisted of measurement of fault plane strike and dip, fault slickenside trend and plunge, and slip direction. The data are tabulated in Table 1. The fault data were then plotted on stereo nets to determine the stress orientations present at the td of faulting (Figure 3). The determined stress orientations are 32~0 included in Table 1. The stereo net plots show average principal stress orientations of: 1) [ ] plunge 1 degree. These data suggest that the N 55 to 70 W oblique-slip faults were produced by a nearly

Jan 1, 1986

By A. Chieregati

"A heterogeneity study was conducted at Erdenet Copper Mine in Erdenet, Mongolia, to determine the fundamental sampling error associated with blasthole sampling. Using a custom-designed radial sector sampler , six large blasthole samples were collected to perform the test developed by Pierre Gy to estimate the constitutional heterogeneity factor for several types of ore. The study investigated the entire sampling protocol, from mine site to analytical laboratory, and quantified the contributions to total fundamental sampling error (FSE) of the various stages of sampling and sample preparation. The total relative standard deviation (RSD) of sampling error was determined to be ±32 percent and ±49 percent for copper and molybdenum, respectively. A classical statistical analysis conducted on the data confirmed the level of RSD native to the mine. About 95 percent of the sampling error was produced from one step in the sampling protocol, that of reducing the 1-kg (2.2-lb) sample to a 50-g (0.1-lb) sample for chemical analysis. To minimize the FSE, the use of a slightly different protocol is recommended. The 1-kg sample (2.2-lb) should first be crushed to a 0.2 cm (0.08 in.) particle size. A 250-g (0.5-lb) split should then be taken and finely ground, from which a 50-g (0.1-lb) split should then be taken. These simple steps will reduce the FSE to ±7.9 percent and ±17 percent for copper and molybdenum, respectively. The modified sampling protocol effectively reduces to zero the probability of misclassification of low-grade ore. IntroductionThe Erdenet copper-molybdenumopenpit mine is located in Erdenet in Orkhon province, Mongolia. The Erdenetiin-Ovoo copper-molybdenum deposit is 1.5 to 3 km (0.9 to 1.8 miles) wide and about 25 km (16 miles) long (Erdenet Mining Corp., 2016). Chalcopyrite, bornite covellite and chalcocite are among the major minerals found in the deposit. The mine processes approximately 26 Mt/a (29 million stpy) of ore."

Jan 11, 2017

By Ted D. Johnson

INTRODUCTION The tailing dam is part of the SOHIO L-Bar Uranium Mine and Mill located about 70.8 kilometers (44 miles) west of Albuquerque in northwestern New Mexico. The dam location is shown on Figure 1. The tailing dam is one of the last upstream method dams permitted in the industry. Design and construction aspects of the project, as well as operation and maintenance considerations, are reviewed herein. Since initial construction, SOHIO Western Mining Company and Woodward-Clyde Consultants have developed several modifications and improvements in the dam&apos;s operating and monitoring procedures. This work has been done in coordination with the New Mexico State Engineer&apos;s Office and in accordance with recommendations in a report done for the Corps of Engineer&apos;s Phase I Dam Inspection Report. Modifications and improvements described should be of particular interest to owners and operators of other tailing dams utilizing the upstream method of deposition. The tailing facility was designed for a 10-year life to provide an ultimate storage capacity of 6 1/2 billion kilograms (7 1/2 million short tons) in an area of about 180 acres. A plan view of the tailing dam is shown on Figure 2. The tailing consists of approximately 75% sand size material and 25% fines (material passing the No. 200 sieve). The tailing dam includes a small earth starter dam with raises of the dam accomplished using tailing deposited by upstream method techniques. The starter dam is a homogeneous embankment about 610 meters (2000 feet) long, a maximum height of 12 meters (40 feet) and 2 1/2:1 (horizontal to vertical) upstream and downstream slopes. The starter dam contains an internal drainage system which includes a near vertical chimney drain connected to horizontal finger drains. A typical section through the maximum height of the starter dam is shown on Figure 3. A small confining dam, approximately 1.6 meters (5 feet) high, is constructed beyond the limits

Jan 1, 1980

By Kim A. Lapakko, David A. Antonson, Jon R. Wagner

This study examined the potential of mixing rotary kiln fines (RK fines), a waste product generated when limestone is converted to lime (CaO), with acid producing mine waste to provide 1) short-term mitigation by neutralizing acid produced by iron sulfide oxidation and 2) long-term mitigation by facilitating the passivation of iron sulfide mineral surfaces. Five loadings of RK fines were mixed with 75 g of Duluth Complex rock to produce neutralization potential (NP) added: acid potential (AP) quotients of 0,0.11,0.22,0.66 and 1.1. The mixtures were subjected to wet-dry cycling laboratory testing for 585 weeks. The rock alone produced acidic drainage, with pH decreasing below 4 after about 20 weeks of reaction and ultimately reaching the approximate range of 3.3 to 3.5; sulfate release did not decrease over time. The RK fines elevated pH and inhibited sulfate release for time periods which increased with the mass of RK fines present. The pH of drainage from the 0.11 quotient (NP added1AP) loading decreased below 6.0 after 75 weeks of dissolution and decreased to 3.6 after 117 weeks, at which time the reactor was terminated. The duplicate 0.22 quotient loadings of RK fines decreased below 6.0 after 95 and 170 weeks. The former was terminated after 117 weeks, and the latter reached a minimum pH of 2.8 after 512 weeks. Drainage pH from the 0.66 quotient decreased below 6 after 581 weeks. For the drainages which acidified, the cumulative calcium release prior to pH decreasing below 6.0 was 92 to 103 percent of that initially present as calcium carbonateloxide in the RK fines. The pH of drainage from the 1.1 quotient typically remained above 8.0 throughout the period of record. Sulfate concentrations in the drainages decreased while the neutralizing fraction of the RK fines was present, but increased after this fraction was depleted.

Jan 1, 2000

By J. D. Miller, C. L. Lin

From mine to mill, the 3D characterization of particles ranging in size from meters to micrometers is now possible. A brief review of X-ray tomography for mineral processing is given with recent advances considered, such as high-speed scanning and image analysis procedures to describe important particle properties, including size, shape and composition. Considering these new 3D characterization tools, applications in mineral processing are discussed for high-resolution X-ray microtomography (HRXMT) laboratory analysis of particles ranging from 20 mm to 5 μm, with a voxel resolution of about 1 μm, including (1) multiphase particle characterization by HRXMT and comparison with scanning electron microscopy (SEM), (2) particle damage state as a result of blasting/comminution, (3) liberation characterization to describe expected separation efficiency, (4) exposed grain surface area analysis to explain the flotation of locked particles, and (5) floc size, shape and water content for polymer-induced flocculation. Finally, applications in mineral processing for plant-site analysis using high-speed X-ray computed tomography (HSXCT) are examined, including, for example, coal washability analysis, the analysis of crusher plant products, and the analysis of pebble phosphate products from central Florida. Scanning rates on the order of 1 kg/min for particle sizes ranging from 150 mm to 1 mm are possible, with a voxel resolution of about 100 μm. For coarse particles smaller than 5 to 10 mm, high scanning rates of more than 300 t/h may be possible at a voxel resolution of about 1 mm.

By Scott A. Norby

Kennecott&apos;s Ray Mines Division (RMD) is located in central Arizona. The pit and silicate plants are located at Ray while the 27,500 tpd concentrating facilities are located at Hayden, approximately 20 miles away from the pit. Ore is delivered in 100 ton railroad cars, 60 cars per train via the Copper Basin Railroad. The side discharge cars are dumped into a 6,000 ton trackhopper. There are 6 Nico feeders, two per belt. Three 48" belts transport the -8" ore to three 7&apos; standard (secondary) Nordberg crushers. Each crusher is preceded by a 6&apos; x 12&apos; Tyler double deck screen. The secondary crushers are open circuit. The -1/21t material from the screens is transported directly to the fine ore bin at the mill. The discharge from the crushers and the intermediate product -2 1/411 x +1/2" is conveyed to a surge bin. The tertiary crusher feeder belts discharge onto five 61 x 161 Tyler double deck screens, each followed by a 7&apos; shorthead (tertiary) Nordberg crusher in closed circuit. The -1/2" product joins the -1/2" ore from the secondary crushers and both are conveyed along a 42", 4,230 foot long belt to the fine ore bin that feeds the grinding circuits. When hard ore is experienced at the concentrator the daily throughput decreases by up to 10%. This coupled with the fact that the ore is generally becoming harder with mine depth provided the incentive to improve the performance at the crushing plant.

Jan 1, 1986

By Saurav Kumar, Akshay Kumar, Vinayaka R. Kiragi, M. K. Banerjee, Amar Patnaik, Vinod Kumar

Today’s industrial technology offers to optimize and reutilize the resources available in terms of scum that remains after hot rolling in steel rerolling mills in Rajasthan, India. Vast reserves of low-grade coal exist and tons of mill scale are produced everyday after hot rolling in steel rerolling mills in Rajasthan, India. The present research work aimed to optimize the various parameters of solid-state reduction of mill scale using low-grade coal from Rajasthan, India. This study investigates the effect of reduction temperature (850–950 °C in steps of 50 °C), reduction time (60 to 180 min, in step of 60 min), mill scale size (800– 300 μm), and mill scale to coal ratio (1:1, 1:2 and 1:3) on achievable metallic iron percentage and its metallization. A higher percentage of iron content is obtained at a lower particle size of 300 μm. The structural analysis is carried out on 300 μm mill scales using XRD analysis, which shows the percentage of iron is approximately 63%, and metallization above 80% can be obtained under the reduction temperature at 900 °C for 180 min, for 1:3 ratio (mill scale to coal ratio). The chemical composition of the mill scale is obtained using X-ray fluorescence (XRF), and particle sizes as well as distribution of mill scale are determined by scanning electron microscopy (SEM). The elements present in the mill scale are also verified using EDS, which indicates the presence of a major portion of Fe, followed by Fe2O3, Si, and C, respectively.

Aug 14, 2020

By Ricardo D. Presnell, Erich U. Petersen, William J. Tafuri, Donald M. Hausen, Russell C. Babcock, William L. Gunter, Thomas R. Gibson, Douglas N. Halbe, Jay W. Hammitt

The Barneys Canyon orebody is a sediment-hosted disseminated gold deposit located 8 km (5 miles) north of the Bingham Canyon porphyry copper mine. Mineable oxide reserves are 8.6 million tonnes (9.5 million short tons) grading 1.64 g/tonne (0.048 oz/ton) gold. Host lithologies are carbonates and siliciclastics of the Park City and Kirkman-Diamond Creek Formations of Lower Permian age. Gold distributions are controlled by favorable stratigraphic units and associated alteration, and structures. Dominant alteration is decalcification and bleaching of carbonates, and argillization and bleaching of sandstones. Gold to silver ratios are greater than 1:1. Anomalous trace metals include arsenic, antimony, mercury and thallium. The Melco deposit lies 2.4 km (1.5 miles) southwest of the Barneys Canyon deposit. It is spatially associated with a steeply northwest-dipping siliciclastic breccia in the Kirkman-Diamond Creek Formation at the contact with underlying orthoquartzites of the Freeman Peak Formation. Both gold-bearing oxidized and sulfide lithologies are present, often in sharp contact. Mineable oxide reserves are 2.8 million tonnes (3.1 million short tons) grading 2.40 g/tonne (0.07 oz/ton) gold. Anomolous trace metals include arsenic, antimony, mercury and thallium, which are generally present in concentrations greater than those at Barneys Canyon, particularly in sulfide-bearing material. A spatial, temporal and genetic relationship is proposed between the Melco deposit and late-stage gold mineralization at Bingham Canyon, although both the Melco and Barneys Canyon deposits have many similarities to Carlin-type gold deposits.

Jan 1, 1990