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By H. T. Mikkelsen

Mineral production implies emissions of particulate matter (dust), noise and vibration that may cause conflicts with neighbors living nearby. In Norway a few conflicts has been brought to court, the latest in 2006. This paper discusses dust emissions and dust depositions (dustfall). The pollution authorities have accepted monitoring of dustfall instead of suspended dust as the first step in conflict cases concerning small and medium sized enterprises. In the case of exceeded limit values, the second step is to require monitoring of suspended dust and next the implementation of control actions to reduce dust emissions. Dustfall monitoring programmes carried out at six mineral production plants in Norway, show that the mineralogical rates of dustfall are rather low a few hundred meters away from the plants or the tailings disposal. They are mostly considerable lower than the limit value of 5 g/m²/30days accepted by the authorities. Seasonal variations between mineralogical and organic dust fractions can be observed in Norway, with higher content of mineral dust during winter time. Other sources to mineralogical dust can be recognized. During winter time wear of road pavement and sand on icy roads can give high emissions of mineralogical dust to the ambient air. Parallel monitoring of suspended dust and rates of dustfall indicate that dustfall can be used to estimate the level of suspended dust. A model for estimating suspended dust based on records of dustfall is presented.

Jan 1, 2007

By J H. Kyle

Cyanide is a significant cost in gold processing, as well as an environmentally sensitive issue. It has been suggested that precipitation of cyanide from leach liquors as copper-iron-cyanide compounds such as Cu2Fe(CN)6 may account for part of the total cyanide consumption. This may be especially true for ores that are rich in both copper and iron minerals. Ammonia addition has been shown to decrease the cyanide consumption by copper. Cu2Fe(CN)6.xNH3 precipitates have also been postulated to form during conventional cyanide detoxification processes, using oxidants such as SO2/air, with ammonia arising as a breakdown product from the oxidation of free cyanide. The present work has examined the conditions under which such precipitates may form, particularly in gold plant process and detoxification liquors. Ammonia and pH were found to have a large effect on the extent of precipitation. The compounds that were formed were found to comprise mainly Cu2Fe(CN)6.7H2O, as well as a range of NH3-bearing variants. The stabilities of the various precipitates are important from an environmental perspective. The long-term stability of copper ferrocyanide precipitates in tailings dams may be a problem under alkaline conditions and when some free cyanide is present, elevating the measured total cyanide in solution significantly. Case studies including data from actual plants are also considered, and the importance of these effects is reaffirmed from these results.

Jan 1, 2000

This excellent paper by Dr. Colombo parallels quite closely the bench and pilot work done by the Hanna Research Laboratory on the Mesabi oxidized taconites. Because of the vast reserves and future ore requirements the need to perfect the selective flocculation-flotation process becomes very necessary. The selective flocculation process is a very efficient method of desliming, leaving a fairly clean, upgraded product for flotation. However, one of the problems that we have encountered in this process was some very fine, -2 micron liberated goethite in the slimes that did not flocculate. The complete slime removal is not necessary. We remove slimes to the extent of about 5% of the iron units. However, on difficult treating ores we would remove more slimes. The reclaimed water from our pilot plant has a turbidity of about 1000-3000 ppm. We can obtain lower turbidity water but to do this would involve additional lime as a tailing flocculant and we would run the risk of flocculating the entire ground pulp in the selective flocculation process because of the increased Ca level in the water. We have about seven different geological ore types out of 17 recognizable ore types that are treatable. All of these ore types require different degrees of dispersion and grinding which is a problem. The use of acrylic acid polymers as a selective flocculant has been patented by the American Cyanamid Co. This will-have to be

Jan 1, 1979

By T Ye Koltunova an, L A. Otrozhdyonnova

This paper presents the results of studies on the development of an efficient non-cyanide technology for processing massive, finely disseminated, copper-zinc ores of the Urals deposits. Copper is mainly represented by chalcopyrite and zinc û by sphalerite. The main ore mineral is pyrite, with its content in ores varying from 50 to 80 per cent. At the present time, separation of copper concentrate at the Urals concentrating plants is achieved in a lime medium by non-cyanide technology using sodium sulfide and zinc sulfite as depressors of zinc and iron sulfides and butyl xanthate as the collector. Copper content in copper concentrates amounts to 15 - 17 per cent; considerable losses of zinc are associated with copper concentrate and final tailings. Various collectors were tested during the studies: xanthates and aerofloats with different lengths of hydrocarbon radicals, thionocarbamates, Aerofine 3418A of Cytec Industries, reagent S-703G of Mineral Reagents Int, and modifiers û sulfur dioxide, sodium sulfide and zinc sulfite, with a pH of the medium varying from 5 to 12. Optimum results of copper concentrate separation were achieved in a weakly acidic medium using S-703G as the collector and sulfur dioxide and zinc sulfite as modifiers. The copper content in the copper concentrate was increased by 4.8 per cent. Total copper recovery into the copper concentrate and zinc recovery into the zinc concentrate exceeded five per cent.

Jan 1, 2005

By J. D. Mooney

(San Francisco sleeting, September, 1915) THE accompanying chart has been drawn for the convenience of engineers as a means of quickly determining the correct number of plies of conveyor belts operating under specific conditions. The calculations are based on the average safe strength (factor of safety, 15) of the various standard rubber conveyor belts. The, calculations assume maximum loading conditions; that is, the belt is considered as carrying the greatest load that it will handle without spillage at ordinary belt speeds. This not only produces the most economical operating conditions, but also the maximum tension in the belt. The chart is a graphical representation of the formula: p = kgW (L + 1OH) Where, p = the correct number of plies k = a constant, depending on the type of drive g = the weight in pounds per cubic foot of material handled W = the width of the belt in inches L = the length of the belt in feet (approximately twice the center distance). H = the difference in elevation between the head and tail pulleys, in feet. For a simple drive, with a bare pulley, k = 1/250,000 For a simple drive, with a rubber-lagged pulley, k = 1/300,000 For a tandem dive, with bare pulleys, k= 1/375,000 For a tandem drive, with rubber-lagged pulleys, k = 1/455,000 The chart is drawn for a simple drive, with a bare pulley (k =1/250,000;

Jan 9, 1915

By Peter J. Stewart

My first experience with carbon reactivation in the gold industry was at Whim Creek, Meekatharra some 15 years ago. Very little was known, in fact, carbon reactivation kilns in the gold room were a complete mystery. Ten years ago at conferences such as this the mystery was more as to why field kilns would not reproduce activities achieved in the laboratory. A carbon reactivation kiln manufac­turer was thought to be brave or stupid to speak at a venue such as this. Reactivation kiln design has improved dramatically in the past 10 years and with a few exceptions, most kilns operate well. May I suggest that now, the focus is on plant practice rather than kiln design, and that the basics of plant practice as it relates to carbon activity needs revisiting. As metallurgists like to talk on engineering, accountants on the Economy and vice versa, my presentation on Carbon Reactivation will focus on plant design and operation as it effects carbon reactivation. There have been some excellent papers published over the past 15 years on the basics of carbon reactivation in relation to gold processingl,2,}. Taken that kilns operate and are effective to a greater or lesser degree, plant design and operating practise have an enormous bearing on carbon activity, solution tail, carbon loss and other cost items. This presentation will not argue the merits of equipment used for the thermal reactivation of organically fouled carbon, questions on this subject are welcome over a drink after this session.

Jan 1, 1995

By Steve Lloyd, Mary Anne Antonioli, Byron Darnton

The Pegasus Gold/Montana Tunnels Mine is a high­tonnage, low-grade, base and precious metal operation in Jefferson City, Montana. Gold heads are typically 0.02 opt with the dominant gold mineral being a native metal electrum quite rich in silver. Early research and development indicated the ore to be amenable to gravity concentration at a 10 to 14 mesh grind. Although full flow gravity preconcentration was considered as a flowsheet alternative, flotation was chosen as the main recovery process due to its apparent lower cost for capital and operating. At this early juncture, the option or concept of recovering coarse free gold by gravity was deemed unnecessary or irrelevant. The intermittent loss of free gold has been observed in the flotation tailing since mill start up. A comprehensive research program was initiated in 1990 to identify the magnitude of this loss and investigate alternatives for its recovery. Batch and pilot plant testing has ultimately led to a permanent gravity circuit being installed at the cyclone underflow. The circuit is designed to scalp out the liberated coarse gold contained in the ball mill circulating load. The final flowsheet consists of: a pinched sluice to upgrade cyclone underflow; a Tyler doubledeck screen to remove over­size from sluice concentrate; Knelson concentrators for roughing and cleaning of screen undersize; an Eriez wet drum separator to remove tramp metal from Knelson concentrate; a Sweco screen to size Knelson cleaner concentrate for tabling; and a Wilfley table to produce final concentrate for direct melting.

Jan 1, 1992

By R. R. Spencer, C. A. Campbell, R. J. Marcotte, A. S. Kromer, P. H. Ostlender

Calumet Div. of Calumet & Hecla Inc. is engaged primarily in mining, milling, and smelting the native copper ores of northern Michigan. The copper occurs in fragmental tops of lava flows and in certain interbedded felsitic conglomerates of Pre- Cambrian age. The most famous native copper ore- body in world history is the Calumet conglomerate, which outcrops between the villages of Calumet and Laurium for a distance of about 2 miles. The Osceola lode parallels the conglomerate lode and is mineralized for about 3 miles. Horizontal distance between the two is 750 ft. In 1932 the price of copper dropped to 54 per lb. Coupled with the high cost of obtaining copper from extreme depths, this low price made the venture un- profitable, and the mining unit was abandoned. After the shaft pillar copper was removed from several of the conglomerate shafts in a retreating operation, the mine was allowed to fill with water. In post-depression years, when the price of copper stabilized at about 12¢, Calumet Div. found it profitable to operate its remaining mines on other lodes, as well as its tailings reclamation plants.

Apr 1, 1956

By D. G. Foot, J. D. McKay, R. A. Ynchausti

The increasing popularity of column flotation as an efficient and profitable mineral beneficiation technique has resulted in the need to perform a study of basic column operating parameters in order to understand mineral behavior in a column. Variation of column rougher flotation parameters on the Nevada Fish Creek fluorite and Utah Collins manganese ores enabled characterization of four beneficiation zones within a column flotation cell. Although specific mineral behavior in each zone was peculiar to the mineral system involved, particle retention time predominantly affected both manganese and fluorite recoveries in the collection zones of the column while enhancing mineral grades in the cleaning zones. Column parameters such as column length, feed injection location, and tailings recirculation velocity were inherently linked to particle retention time and retention time distribution, both of which affected mineral recovery. Froth depth had a discernible effect on fluorite concentrate grades; as froth depth increased, fluorite grades increased. Wash water additions affected both fluorite and manganese recoveries and fluorite grade. Axial mixing drastically reduced fluorite and manganese recoveries due to short circuiting of some of the feed slurry in the column.

Jan 1, 1988

By Costello MC

An alternative process for desorbing and recovering metals from activated carbon has been developed, and is the subject of several patent applications by Micron Research (W.A.) The system uses the principle of conditioning the carbon in a strong alkaline cyanide solution followed by distillation desorption in a fractionating column, using the carbon as the column packing and methanol as the desorbing agent. The system offers the following advantages : (a) The processes of desorption, methanol recovery and reactivation are all carried out in a single column with the carbon itself acting as the column packing. No other desorption plant items or fractional distillation equipment are required. (b) The metal values are concentrated into the water content of the carbon and no other water is used except as condenser coolant. This produces very high grade gold silver solutions which give high current efficiencies in the electrowinning step. These high gold concentrations allow almost complete separation of gold and silver from copper in the electrolytic cell. (c) The process is completed to a tailing assay of less than 200g Au/tonne carbon in 15 - 20 hours allowing turn around of carbon batches on a daily Basis.

Jan 1, 1982

By W. H. Eddy

The Bureau of Mines conducted laboratory batch-and small-scale continuous flotation processing tests on waste granite fines from Georgia, to determine the feasibility of recovering usable feldspar and quartz products by flotation. Two flotation methods were utilized for the removal of biotite and other iron-bearing minerals: One was in an acid-cationic circuit, and the other in an alkaline-anionic-cationic circuit. Feldspar was recovered using an acid-cationic method. To produce commercial-grade feldspar and quartz products it was necessary to employ either two sequential froth flotation circuits or one froth flotation circuit supplemented by wet-magnetic separation techniques to remove iron contaminants. Continuous flotation processing yielded a commercial-grade feldspar concentrate containing, in percent, 3.8 Na20, 5.4 K20, and 1.10 CaO, with recoveries of 77.8, 70.3, and 64.4, respectively. The feldspar tailing was high-quality glass sand containing 98.2 percent SiO2 and 0.024 percent Fe2O3. About 42 percent of the quartz was recovered at glass sand specifications. An effective separation of, potash feldspar from the bulk feldspar concentrate was achieved by batch flotation in a mineral separation cell.

Jan 1, 1974

By Lothar Anheuser

Theoretical research, laboratory tests and practical experience confirm the possibility to stabilise the excavated tunnel front by suitably composed slurries even in coarse ground with extremely high permeability. The fluid assures dependable repartition of the pressure, low torque of the cutting wheel and small perturbances of the original ground structure. Very easy access to the front, for tool repair or removal of obstacles, are other features of the WF-Hydroshield concept. For soft ground, the Hydrojet (excavation by immersed slurry jets) permits still further improvements. Field experiences in very mixed face conditions are related. Measured soil subsidences in homogeneous, loose sands above Hydroshield-driven tunnels are compared to calculated values. Dependable shield tail seals, which are a necessary condition for work without compressed air in the tunnel, and an original repair kit are presented. The aim to liberate shield driving from hazards of narrowly defined soil conditions and to broaden the range of applications led to a series of bentonite shield machines, derivated from the initial slurry type, such as Dense Mud Shield, Earth Pressure Balance Shield or shields with hard rock cutterheads. Similar to these constructions, overall stability of the front was the major concern in developing the Hydroshield. Initially, this type was designed for use in cohesionless medium to coarse sands and gravels only. The mechanism of front collapse, initiated by a local slide uncovering a limited area, which triggers a

Jan 1, 1983

By K. R. Weller

The Ardlethan Tin Ltd. concentrator typically recovers 55% of the tin in a gravity concentrate at 51% grade from fresh ore assaying 0.5% Sn. Coarse gravity concentration is performed in a complex circuit involving Reichert cones, spirals, shaking tables and two regrind mills. The paper describes how flow and size-by-size sample data from a steady-state survey were used to construct a computer model of the plant involving 46 process streams. The data reduction involved a computer- aided mass balancing technique, recovery-size modelling of gravity concentrators and the use of a mineral dressing flowsheet simulator. The model was used to simulate possible rearrangements of the available equipment. Early simulations indicated that improved performance could be obtained by using spirals to-treat cone middlings or tailings rather than cone concentrates. A test spiral rig was used at Ardlethan to determine typical performances for all cone product streams and the results were incorporated in new simulations. A revised circuit was recommended and implemented by Ardlethan Tin as far as physical restraints within the plant would allow. Production figures show that a substantial increase in tin recovery and 10-15% increase in milling rate were achieved following the circuit changes, and these improvements were sustained over a six month period.

Jan 1, 1984

By W. Blaschke, E. Malysa

INTRODUCTION In recent years in Poland, a problem connected with the occurrence of a considerable amount of liberated galena grains in size fractions under 1mm has arisen in the zinc-lead processing area. Operating practices have included crushing of ore particles under 0.2 mm followed by flotation. Jigs were applied to separate the concentrates from size fractions under 10 mm and to prevent the galena overgrinding. However, the fractions being too wide, the jigs have not worked effectively, in spite of the great differences in specific gravities of galena, blende and gangue minerals. Fine galena grains move into light fractions in considerable amounts and hence are overground. Therefore, the flotation becomes difficult and galena is lost in its tailings. These losses could be avoided by introducing separation in the grinding and classification circuits. The amount of 0-1 mm fraction is great; therefore, the application of concentrating tables or other kinds of jigs is impossible from an economic point of view. Taking these factors into account, the possibility of application of pinched-sluices was investigated. These are simple devices without moving parts, they have a high output for the unit areas, and their water and energy consumptions are low. In this paper, the results of zinc-lead ores beneficiation in pinched-sluices are presented.

Jan 1, 1980

By D. W. Hutchinson, W. W. Wen, A. A. Terchick, J. C. Anderson

It is estimated that approximately 3% of the raw coal processed today in coal preparation plants throughout the nation will ultimately report to slurry-tailing disposal (Anderson, 1975). The usual disposal method is to pump refuse slurry containing - 48 mesh (300 pm) refuse into dams or impoundments constructed of coarse refuse. The failure of an impoundment at Buffalo Creek, Saunders, WV (Anon., 1972) resulted in an evaluation of refuse disposal methods by various regulatory agencies as well as coal producers. The result of this evaluation was the promulgation of federal regulations dealing with the construction, operation, and final abandonment of coal refuse disposal area. As a consequence of these regulations, the use of the landfill/impoundment method of disposal became in some cases economically unfeasible or technically impossible. US Steel has studied a number of alternative methods of refuse disposal. During an investigation of mechanical dewatering with a pressure filter, portland cement was added to the slurry to solidify the filter cake. Because of the success of the cement addition, it was decided to study the stabilization of fine coal refuse slurry through use of cement-type additives. The goal of the investigation is to develop a process that will use cement-type additives to solidify the refuse slurry into a form that can be handled and stacked with the coarse refuse, thus eliminating the need for slurry impoundments.

Jan 1, 1982

Silica-Kaolin Sand Wet Sizing The Bulgarian deposits of silica-kaolin sand are localised in the eastern part of the Danube plain. After wet separation the raw material yields two products - washed kaolin (different grades) and silica sand (different grades). The water washing technology includes: disintegration; spiral classification; hydrocyclone classification and dewatering of kaolin grades; hydroclassification; sizing using drum screens and dewatering of silica sand grades. Research and development program has started to improve the silica-kaolin sand separation in processing plants. Based on the raw material and tails characterisation data and the preliminary experiments using Derrick Stacksizer screening machines a silica-kaolin sand wet sizing technology was developed. The full-scale pilot flowsheet includes the following units: Derrick Stacksizer equipped with 0.43 mm opening Derrick Urethane Panel; 500 mm Flat bottom Hydrocyclone; Derrick Stacksizer equipped with 0.150 mm opening Derrick Urethane Panel; 100 mm Hydrocyclone and Derrick Dual Motor Linear Motion Machine equipped with 0.064 mm opening Derrick Pyramid Panel. The pilot-test data clearly demonstrate that the new silica-kaolin sand wet sizing technology is suitable for Bulgarian silica-kaolin raw material. There are possibilities to increase the silica sand recovery increasing the production of silica sand grade PK 0.16 and starting the production a new grade PK 0.10.

Sep 13, 2010

By S. K. Kawatra, T. C. Eisele

Coal-cleaning operations commonly produce a fine-waste stream containing a considerable amount of coal. The recovery of this coal is normally not economical because the coal is mixed with large amounts of difficult-to remove clay slimes and is often heavily oxidized or contaminated. Conventional froth-flotation machines cannot, in a single flotation stage, produce an acceptable product from such material, and even flotation columns often have difficulty cleaning this type of feed. A new type of flotation column was designed specifically to deal with these problems, and a column of this design was installed and tested in a coal-cleaning plant in Ohio. The feed coal tested was the plant&apos;s tailings-thickener product. The coal was 80% -176 µm and had a heating value of 8401 Btu/ lb. The coal contained 39.8% ash, as fine clay, and 2.83% total sulfur. Because previous attempts at conventional flotation of this type of material had been abandoned, plant personnel believed that the coal was too heavily oxidized and contaminated for flotation to be effective. However, by using the new column it was possible to recover up to 85% of the calorific value in a product containing < 11 % ash, and it was possible to recover 60% of the calorific value in a product containing <8% ash. In addition, the bulk of the pyritic sulfur was also rejected from the coat.

Jan 1, 1996

By S Cossfo Torres

Minera PenmontÆs gold mine, La Herradura, is located at the northwestern Mexican state of Sonora and consists of an open pit mine and heap leaching facility for the recovery of gold. The production began in July 1998, and today La Herradura produces about 200 000 oz of gold per year. Standard sampling and assaying protocols were established during exploration based on heterogeneity tests. However, at deeper zones of the pit, coarse gold became evident and it was necessary to develop a new sampling protocol in order to minimise the fundamental error and improve the metallurgical accounting. This new sampling protocol requires cyanide leaching the whole pulverised sample from the blasthole, assaying pregnant solution and fire assaying 50 g of tailings. A sampling study was performed on 195 blasthole samples analysing gold under the standard and new protocols. The new protocol demonstrated that the original one underestimated gold grade by approximately 27 per cent, with impacts on ore misclassification and metallurgical accounting. According to the ore grade control system, mineral with gold grade above 0.35 g/t was sent to the leach pad and mineral with lower grade was classified as waste. Based on the sampling study, it was decided to classify a low-grade ore with gold assay between 0.20 and 0.35 g/t which was stored for future treatment under better gold prices or should mine production decrease. These two conditions were present during 2005 and 2006, when about 3.5 Mt of low-grade ore were leached.

Jan 1, 2008

By F. Lincoln & B. Rodier, G. Poniatowski, N. Kawashima, T. J. White, J. H. Kyle, G. F. Eaton

The history of Australia is a narrative of the discovery and development of its mineral wealth. The gold strike at Ballarat, Victoria, in 1851 transformed the colony by triggering the first great influx of migrants to Australia, and the development of subsequent mineral discoveries lead to a vast resource industry. However, beneficiation of base and precious metal ores carried with it an environmental penalty, not addressed until the last quarter of this century. In particular, the extensive deposits of gold, silver, lead, copper and zinc reside in sulphurous ores which often contain arsenic, antimony and mercury as significant contaminants that are released as unwanted by-products during smelting and refining. Until recently, these toxic concentrates were discharged to the atmosphere or receiving waters, collected in tailings ponds close to the smelters, or dumped at sea. Although such procedures . were acceptable in the past, more stringent environmental controls have required a re-examination of these practices and encouraged intensive efforts to develop enduring solutions to the disposal of these wastes. From this research has emerged a promising new generic technology called Xtaltite (pronounced crystal-tight). In this approach, geochemical principles are employed, first, to reduce the toxicity of heavy metals to low levels, and second, to identify and engineer suitable repositories for the permanent disposal of the consolidated waste form. In this paper, we emphasise the application of Xtaltite to the disposal of arsenious smelter fumes derived from gold winning. The principles ofXtaltite will be described and its technical and economic performance compared with conventional technolo­gies.

Jan 1, 1995

By Vernom F. Swanson, George E. Fish

A study was made during 1959-60 of reported nelsonite occurrences in Nelson and Amherst Counties, Va. The investigation was limited to core drill- ing on one deposit and beneficiation studies on three deposits. Three diamond-drill holes totaling 562 feet were drilled on the E. M. Hughes property to determine subsurface continuity. Although outcrops as much as 30 feet wide were present on this property, the only nelsonite intersected in the holes ranged from 4 to 18 inches thick. The ilmenite content of the core ranged from 50 to 75 percent. These nelsonite deposits are of secondary importance as a source of titanium, being small, rarely exceeding 1,500 feet in length and 30 feet in width, but they are high grade. Due to the proximity of some of the bodies it might be possible to mine several of them as one operation feeding one mill. Electromagnetic separation on composited samples recovered 83.0 percent of the Tio as an ilmenite concentrate in 68.5 percent of the weight. Two flotation procedures were developed. The first method, using oleic acid as a collector and fluo silicic acid as an apatite depressant on deslimed feed, produced an overall recovery of 59.3 percent in batch tests. The second procedure eliminated desliming. Laurylamine hydrochloride removed silicates in an acid pulp followed by an alkaline apatite flotation with oleic acid as the collector and dextrine as the ilmenite depressant. The ilmenite-rich tailing represented a 66.2-percent recovery of Tio in the batch tests.

Jan 1, 1964