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By S. K. Pan, M. P. Srivastava, N. Prasad

"Dalli mines of Bhilai Steel Plant, India is fully mechanized iron ore mine having wet processing system. It has a capacity to treat 5 Mt of ROM. In processing circuit of Dalli mines, scrubbing, wet screening and classification operation is being carried out. Average slime generation for the last five years is about 14% of ROM. As the mining has reached to the lower benches, fines generation is more. As a result, very fine and heavy particles, which settle fast, are being discarded with the slime to tailing pond. These heavy and faster settling particles were causing jamming in the thickeners.Laboratory investigations were carried out at Research and Development Centre for Iron and Steel (RDCIS) to suggest a suitable scheme to recover high quality fines so that the availability of thickener can improve. The slime samples were first treated in a floatex density separator. The concentrate obtained from the floatex density separator was treated in hydrocyclone. Laboratory investigations indicated that the slime with 49% Fe, 12.7% silica and 10.7% alumina can be upgraded to 62.1% Fe, 3.3% silica and 2% alumina with 36% material recovery. Based on the laboratory studies ,a system was installed at Dalli minesThe general arrangement of the system is shown in Fig. below."

Jan 1, 2003

By Frank Firmstone

THE indicator cards shown herewith were taken by the writer in June, 1877, from the water-pressure blowing engine of the Longdale Iron Co., at Lucy Selina Furnace, Longdale, Virginia. A description of this engine with figures, and some indicator cards, was published by Washington Jones, Esq., Superintendent of the I. P. Morris Co.,* in the Journal of the Franklin Institute for May, 1877, to which those desiring a full account are referred, but for convenience the principal dimensions, etc., are given below. The machine consists of a pair of horizontal directacting blowing engines, connected at right angles to a common fly-wheel shaft. The blowing cylinders are 4 ft. diameter by 5 ft. stroke; the water cylinders were 18 in. diameter by 5 ft. stroke as built. Both water and blowing cylinders are double acting. The water valves are balanced piston valves, 18 in. diameter, moved directly by eccentrics on the fly-wheel- shaft. Great care was taken to have the areas of the ports and passages at least equal to the area of the working piston. The engine is set about 14 ft. above the tail-race, and the * The engine was built by the I. P. Morris Go., from drawings made by Messrs. Tuws and Hartman.

Jan 1, 1879

During the past decade geopolymer technology has emerged as a possible solution for the effective stabilisation and solidification of waste materials. Amorphous to semi-crystalline geopolymers form by the co-polymerisation of individual alumino and silicate species, which originate from the dissolution of silicon and aluminium containing source materials at a high pH in the presence of soluble alkali metal silicates. It has been shown previously that geopolymerisation can transform a wide range of waste alumino-silicate materials such as fly ash and blast furnace slag into building and mining products with excellent chemical and physical properties, including high compressive strength as well as fire and acid resistance. These chemical and physical properties suggest that geopolymeric matrices derived from waste materials are well suited for environmental and other applications such as the encapsulation of mine tails, the immobilisation of heavy metals and also the fabrication of structural products. The current challenge is to demonstrate that a wide variety of natural Al-Si minerals could serve as potential source materials for the synthesis of geopolymers. Moreover, it should be shown on a large commercial scale that it is possible to encapsulate waste materials into the geopolymeric structure without compromising the chemical and physical properties of the final product for the desired application.

Jan 1, 2000

The efficient development and exploitation of an orebody relies on an understanding of the mineral characteristics of the deposit. This information is valuable throughout the mining cycle - during exploration, assessment, testing, design, operation and rehabilitation. Problems can arise at each stage if there is an inadequate appreciation of the mineralogy involved and its effects on the treatment process. Modern image analysis systems now permit a more quantitative understanding of the mineralogy of an ore to be gathered. Where previously the metallurgist had to rely on qualitative information to guide technical decisions, it is now possible to give numerical estimates to the compositional and textural properties which govern the mineral concentration processes. Ore samples can be characterised according to modal abundances, mineral grain size parameters and mineral associations, all of which affect the liberation and concentration of the values. The paper will discuss the application of quantitative process mineralogy by considering a number of case studies. These include: ore characterisation to assist in exploration, assessment and mine planning; modal analysis of complex mineral suites to assist in the evaluation of ores, the deportment of values between products and the recovery or loss of these values; liberation analysis for ore testing, plant optimisation and performance monitoring; and characterisation of tailing to guide estimates of leachability, stability and rehabilitation options.

Jan 1, 1998

By De Cuyper J

Laboratory testwork has been performed on ancient residues from the cyanidation of a previously roasted gold-containing pyrite and arsenopyrite ore. By grinding to 95 per cent minus 60 microns, and after a strong aeration conditioning of the pulp, a gold recovery of approximately 67 per cent was obtained by cyanidation. This recovery was brought up to 80 per cent by leaching with thiourea, after a preleaching step with sulphuric acid under severe conditions of acidity and temperature. To avoid excessive capital costs bound to such processes, it was decided to investigate the possibility of applying flotation after grinding at a size similar or finer than the one used for the leaching tests. The method turned out to give interesting results because of the natural flotability observed at pH values lower than two for particles rich in gold present in the residue. After two cleaning steps, a final concentrate, representing around one weight per cent of the feed, was obtained with an approximate gold content of 350 g/t, representing a batch recovery of gold of about 50 per cent. Apart from sulphuric acid, whose addition was very high (flotation slurries at 25 g/l H2SO4), other reagents were not found necessary. Additional flotation tests were carried out to examine whether the recycling of the cleaner tails was justified but the gain observed appeared insufficient.

Jan 1, 1993

By E. V. Manlapig, M. C. Harris

"The complexity of the flotation process has made the identification of an appropriate model structure for the predictive simulation of flotation circuits far from obvious. Work performed under the auspices of the AMIRA P9 flotation program has indicated that the metallurgical performance achieved from a flotation cell arises from the interaction of two essentially independent property classes, namely the cell operating conditions and the properties of the cell feed stream. These observations have led to the development of a model structure that associates a constant set of flotation parameters to the streams of the circuit, which are transformed into a flotation response depending on the operating conditions within the flotation cell.A variety of alternative model structures exist within the literature which do not make this distinction between cell and stream flotation properties, and many of these models are in widespread use for flotation optimisation and design studies, both in industry and academia.In this paper, circuit survey and laboratory batch flotation data collected from an industrial flotation circuit are analysed. The data demonstrate that streams in an industrial flotation circuit exhibit varying particle floatability – concentrate streams typically float faster than tailing streams (Figure 1(a)). This floatability, however, is generally conserved around the various units within the flotation process (i.e. the effective floatability of the combined feed and product streams of a unit is equivalent (Figure 1(b)))."

Jan 1, 2003

The high-grade orebody at Miami was mined successively by top-slicing, shrinkage, stoping and under caving. The method described in this paper was developed to enable the low-grade orebody (36,000,000 tons assaying 1.06 per cent copper) to be mined at a low cost without further dilution of the grade. The method is clearly described and illustrated by drawings and photographs, and the original paper should be consulted as it is impracticable to abstract it, beyond noting that the mining cost is just under 40 cents per ton and the average grade of ore now being produced yields less copper than the mill tailing of 1915, the tons mined per day having increased from 4000 to 18,000 in the same period. Vertical and Incline Shaft Sinking at North Star Mine By Arthur B. Foote, Grass Valley, Calif. (Tech. Pub. No. 324: Class A. Metal Mining, No. 36) THE main shaft of the North Star mine is an inclined one following the main vein at an average dip of 26 deg., a vertical shaft intersects it at the 4000-ft. (incline) level, the incline continuing to the 6300-ft. level. At this point the vein was found to give out and a new vein, called the No. 1, was found to dip in exactly the opposite direction.

Jan 1, 1930

By James Malcolmson

(New York Meeting, February, 1913.) THE Tigre Mining Co. of Mexico, owned by the Lucky Tiger Combination Gold Mining Co. of Kansas City, decided early in 1910 to enlarge its mill, which consisted of a concentrator milling 3,000 tons of ore per month. The enlarged plant was designed to crush and concentrate 6,000 of ore and to cyanide 7,500 tons per mouth. The extra tonnage cyanided was to be obtained from tailings which were stored below the old mill. Power was supplied to the old mill from four boilers; a 150-h-p.. Corliss engine and two Weber 60-h-p. gas-engines using producer-gas made from charcoal. The total horse-power required, including 60 h-p. used for machine-drills in the mine, was approximately 275. Cord-wood cost $4.32 per cord, 8 by 4 by 3 ft., and charcoal $15 per ton. The fuel supply in the surrounding country was approaching exhaustion; more than 150 mules and 50 men were engaged in cutting and transporting this fuel, and the unsettled political, conditions in Sonora were such that it would be practically impossible to keep an organization of this kind together. The price of fuel was steadily rising and at times it became almost impossible to get sufficient fuel to operate.

Jan 4, 1913

By Walter E. Heinrichs

This case history really started during the tail end of WWII when two Navy men, both Colorado School of Mines graduates, got together in Washington, DC, while they were assigned to the Office of Navy Petroleum Reserves doing some of the first aerial magnetometer work ever done on a practical oil exploration basis over the offshore area of the Gulf Coast. When separated from Naval Service active duty in 1946, the two men, Walter E. Heinrichs, Jr., and Robert E. Thurmond, tried, unsuccessfully, to form a commercial air magnetometer business. Later they got together again in Jerome, AZ, with Newmont Mining's early geophysical exploration endeavors from 1947 to 1949. Subsequent to the Newmont experience, they jointly proposed a specific mining geophysical exploration and development program to Herbert Hoover, Jr., then president of United Geophysical Co. of Pasadena, CA. Mr. Hoover, in consultation with his father Herbert Hoover and others of his staff at United, particularly Edwin G. Schempf, executive vice president, and Raymond A. Peterson, vice president of research, agreed to establish the proposed program under United's auspices, to be headquartered at Tucson, AZ. The primary objective would be to practically test and develop various mining geophysical methods and techniques while exploring particularly for base metals in the general area of southeast Arizona and southwest New Mexico.

Jan 1, 1991

By A Crawford, P Manton

Telfer is a gold/copper operation, currently operating one open pit mine and one underground mine. The ore processing plant consists of two parallel trains involving semi-autogenous mill, ball mill and crusher (SABC) grinding, flotation and leaching. Due to receiving relatively high grade underground ore for over half of its total feed, Train 1 frequently experienced cleaning circuit overloading during high production periods. This resulted in losses in recovery and ineffective concentrate upgrade. As a quick and simple solution, the Train 1 copper flotation circuit was reconfigured to send concentrate from the first two rougher cells directly to the final concentrate and return the cleaner scavenger tailings to the third rougher cell. These modifications have resulted in a 1.4 per cent increase in plant copper recovery with no change in final concentrate grade. Preliminary results also indicate gold recovery has increased by more than 1 per cent since the circuit reconfiguration. Other benefits include less spillage, more friendly circuit to operate and a more consistent metallurgical performance. Focus for the next phase of optimisation is on concentrate upgrade. A number of opportunities have been identified and are currently being investigated, including staged reagent addition, single stage cleaning of the first two copper rougher concentrates, regrinding of the remaining copper rougher concentrate prior to cleaning and process control based on a machine vision system.

Jan 1, 2009

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