The problem involved in Manganese Ore Processing deals with the production of acceptable specification grades of manganese concentrates at a maximum recovery of the total manganese from ores having variable characteristics. The flowsheet provides for both gravity and flotation with a maximum recovery of the manganese values in a coarse size in the most economical manner by the use of jigs and tables. The coarse concentrate must be up to grade and is immediately acceptable to the steel industry. The fine concentrate produced by flotation is made available for nodulizing or sintering.
The present world situation and lack of high grade manganese ores in the western world has had a pronounced influence on the development and utilization of the lower grade manganese ores. The specification stipulated by the Federal Stockpiling program for manganese ores or concentrates requires a fairly high manganese content with minimum quantities of impurities.
The flowsheet incorporates a conventional multistage crushing plant with a grizzly or screen ahead ofboth the primary and secondary crushers. The mine run ore is dumped through a 10 grizzly into a coarse ore bin. The ore is discharged by a Apron Feeder to feed the primary Jaw Crusher. This crusher is equipped with a 2 opening shaking grizzly to remove the undersize material.
The secondary cone crusher is fed with the oversize product from a 3 x 6 Vibrating Screen. This is an example of standard practice of removing all particles as soon as they are reduced to the proper size at each crushing stage. This is important in order to prevent the production of excess fines so easily produced in crushing manganese ores.
Sampling at this point is done by means of Samplers. They cut an accurate sample and are inexpensive to operate and maintain. The material cut by the initial sampler is fed at a constant rate by means of a vibratory feeder to a set of rolls for further crushing prior to the final sample cut. This results in the most accurate sample possible.
Separate bins are provided to temporarily store the ore until the assays on each lot of ore are known. The mill feed can then be drawn from these bins for proper blending of various types and grades of ore as desired. Ore of different types and grades can also be drawn from these bins for stockpiling a supply of blended ore to provide a uniform ore for continuous mill operation.
The crushing and sampling plant is designed to operate on a one shift per day basis with a capacity of from 400 to 500 tons per shift. The excess crushing capacity is to allow for the stockpiling of excess available ore and to take care of the operation on one shift.
The mill feed, drawn from one or more bins, is sampled at the ore feeder discharge to obtain a composite sample for mill control. After elevating, a vibrating screen separates the feed into sizes best suited for the Improved Harz Type Jigs and Selective Mineral Jigs. The coarsest part of the feed goes to the Harz Type Jigs which produces a final concentrate and a tailing. The finer portion of the feed, usually -8 or -10 mesh passes to the Mineral Jig for the recovery of a final concentrate.
The tailings from the Harz Jig are ground in a Steel Head Rod Mill after being dewatered by means of a Crossflow Classifier. The rod grate type mill, equipped with a 10 mesh spiral screen, grinds the jig tailings to minus 10 mesh with a minimum quantity of slimes. The spiral screen removes any plus 10 mesh material which is returned to the classifier. The minus 10 mesh rod mill discharge is combined with the tailings from the Mineral Jig and are pumped to a Hydraulic Classifier for size separation for table fed. Each gravity concentration table treats a separate size range which allows most efficient results. The tables produce a final concentrate, a middling product, which is returned to the rod mill for further grinding, and a sand tailing. The table tailings are either further treated by flotation after regrinding, or are discarded, depending on the assay.
The jig and table circuit can save from 50 to 80% of the manganese, depending on the characteristics of the ore. The grade of the jig and table concentrates is from 44 to 46% metallic manganese. It is essential to recover as much manganese as possible in the gravity concentration section since its milling cost is muchlower than in the flotation treatment, and the simple operation is more positive. This demonstrates the principle of when mineral is free, remove it which is still good metallurgy. Some ores, however, can only be treated by flotation to a greater extent in order to make an overall economic recovery.
Types 1 and 2 ores require a prefloat treatmentto remove the calcite as a froth. The calcite must be removed ahead of the manganese since if left in the circuit it will float with the manganese, thereby giving a low grade manganese concentrate. Tailings from the calcite prefloat circuit are then further treated by flotation, floating the manganese as a concentrate.
Careful and complete conditioning is a very important step in manganese flotation. Here we use a Special Super-Agitator and Conditioner for the proper mixing of the reagents into the pulp plus Super Rougher cells as conditioners. This provides the intense mixing for proper flocculation so essential for manganese flotation. The amount of aeration is easily controlled during the conditioning.
A nodulizing or sintering step may be necessary for the further treatment of the flotation concentrates. This step produces nodules or a sinter product acceptable to industry and the grade of manganese is also materially increased by such treatment.
This flowsheet is designed to produce a maximum amount of the manganese in a coarse form which will be marketable without the further and high cost of nodulizing or sintering. The gravity concentration sections do this. Since the reagent costs for manganese flotation are high and in direct proportion to the amount of flotation concentrates produced, preceding flotation by gravity concentration results in maximum recovery with lowest cost.This flowsheet follows the fundamental rule of metallurgyrecover your mineral as soon as free and as coarse as possible.
As the problem involved stock piling of the minus 20-mesh material for selective flotation recovery under more favorable market conditions, the equipment selected at this stage consisted only of gravity concentration and sizing equipment to produce a partially- concentrated product which could be economically shipped to the purchaser.Atypical manganese oxide ore stipulated that it contain not more than 10% minus 20-mesh material.
Mine ore is dumped through an 8 Grizzly into a coarse ore bin provided with a rack and pinion gate for discharging the ore to the Apron Ore Feeder which is built to resist high abrasion and the stress of sudden impact. A feeder with 30 wide flights was chosen in this case and a sufficient length was recommended for the feeder to allow for a portion of it being used as a picking belt. The availability of low-cost labor makes it possible to discard considerable waste rock at this point.
Primary sizing is done by means of a 3x 5 Grizzly with 2 openings. This Grizzly could be made into the vibrating type if desired, obtaining its motion from the pitman of the crusher. Grizzly undersize passes to a conveyor and oversize to the primary crusher.
The single deck, 3x 6 Vibrating Screen removes the minus 3/8 product from the secondary crusher feed. The minus 2 plus 3/8 product is fed to the secondary crusher, the minus 3/8 screen undersize becoming part of the feed to the jigs.
Several excellent gyratory crushers are on the market. A 1-8 Traylor Gyratory Crusher unit was selected to reduce the minus 2 plus 3/8 feed to all minus 3/8- At this point in the flowsheet it would be possible to utilize crushing rolls which tend to produce slightly less fines than a gyratory crusher. However, due to the greater reduction ratio of the crusher and the difficulty in transporting crushing rolls to the millsite, the gyratory crusher was recommended.
Ratios of concentration as high as 97,000 to 1 have been made Ina Selective Mineral Jig. Far continuous discharge of preciousmetal concentrates, Dowsett Valves with locking arrangementmay be used on hutch discharges.
Two Duplex Selective Mineral Jigs concentrate the minus 6-mesh manganese ore. Tailings from these jigs are sent to waste. The high-grade product produced by the jigs selective action is sent to further screening.
As the market requires that not more than 10% of the shipping ore is minus 20-mesh, the selective jig concentrates are passed over a single-deck, 2x 4 Dillon Vibrating Screen, with 20-mesh screen cloth. The plus 20-mesh screen oversize becomes shipping ore and the minus 20-mesh manganese is stock piled for future marketing. Present briquetting costs do not permit this method of preparation for market at this time.
The minus 3/8 undersize from Screen No. 1, together with the minus 3/8 plus 6-mesh product from Screen No. 2 are concentrated by two 3-compartment, (Improved Harz Type) Jigs. Units with 3 compartments were chosen to give ample capacity to produce a high-grade manganese product. Tailings from these jigs go to waste and the concentrates become shipping ore.
This flowsheet is based upon the principle of recovering the mineral as soon as it is free from the gangue. This is essential in the treatment of manganese ores due to their tendency to slime readily. Note that both the motor horsepower provided for each machine and the actual horsepower required is shown. The motor horsepower figures are enclosed in circles and the horsepower-consumed figures are underlined.
The ordinary specifications for marketing manganese ore are as follows (dry ore basis): Mn, minimum.48.0 per cent Fe, maximum..6.0 per cent P, maximum..0.12 per cent Si02 + Al2O3, maximum11.00 per cent Non-ferrous impurities, maximum.1.00 per cent Size analysis shall show all minus 1 inch and not more than 25 per cent to pass a 20 mesh screen.
While managanese ore is not a non-metallic, the application of flotation to its beneficiation is similar to that used for the non-metallic ores. Non-metallic reagents are used to float non-metallic impurity minerals such ascalcite, and other non-metallic reagents can be used to concentrate the manganese mineral and reject silica and alumina minerals as a tailing. Manganese is a critical mineral in America and the development of new methods of beneficiation is highly desirable for our national defenses. While much investigational work has been carried out by the U.S. Bureau of Mines and others, there is still a need for more efficient reagents to make many ores economically amenable to the flotation process.
(1) Carbonate-Gangue OresThe carbonate gangue, such as calcite, is floated first with fatty acid, usingan alkaline pulp and a starch or yellow dextrine to inhibit the manganese oxide. The pulp is then acidified and the manganese oxide floated with an emulsion of crude Tall oil, and heavy fuel oil emulsified in hot water with petroleum acids such as Oronite wetting agent S or Oronite sulfonate L.
The crushing and grinding operations are an important part of the processing of mineral resources, and it is also operation with high investment and high energy consumption. In the case of metal mines, equipment investment in crushing operations accounts for 65% to 70% of the total plant value, power consumption is about 50% to 65%, and steel consumption is as high as 50%.
Therefore, how to improve the performance of grinding equipment, research and development of high-efficiency energy-saving equipment, obtain a larger crushing ratio, achieve a finer crushed product granularity, reduce steel consumption, has become a common goal pursued by workers in various fields.
The ore size reduction process involves two steps: crushing and grinding. The grinding process is the final operation of making the mineral to dissociate from the monomer and making the particle size meet the selection requirements. Grinding is a high-efficiency and low-efficiency operation. The power consumption of crushing operations only accounts for 8% to 12% of grinding operations. Improving the grinding process is an effective way to achieve high efficiency, low consumption and increase economic benefits.
The crushing of materials is mainly achieved by the extrusion and impact of the equipment on the minerals, and the grinding is mainly achieved by the impact, grinding and grinding of the equipment. The energy utilization efficiency of the crushing operation is much higher than that of the grinding operation. More crushing and less grinding to achieve the best economic benefits.
-Adopt high-efficiency fine crusher. Such as the single-cylinder hydraulic cone crusher, the HP series cone crusher ( Nordberg ), and the domestic JC56, JC4060 jaw crusher, SX series double-roller jaw crusher, etc.
The principle of mineral grinding-classification processing is to combine grinding and classification operation, cleans out the gangue minerals in time, reduces the grinding volume and improves the beneficiation efficiency.
The crushing operation of the ore dressing plant is very inefficient, and the fine crusher replaces the conventional mill to produce fine products. For the hard rock crushing, the water punching cone crusher can gradually replace the conventional drum mill.
Some of the original mining plants have a large design scale, but for a variety of reasons, the production scale is only about half of its original design. With the gradual reduction of mineral resources, these old factories can be improved in energy conservation and efficiency, perfecting their crushing process, ensuring their crushing granularity while achieving energy saving and efficiency.
In the current mining production, the crushing method of mineral materials is mainly mechanical crushing. In order to reduce the steel consumption of crushing operations and improve the efficiency of energy utilization, mining workers have developed a new method of crushing, in which microwave pretreatment is a promising method of crushing.
The microwave is an electromagnetic wave having a frequency of approximately 300 MHz to 300 GHz and a wavelength of 2500 px to 1 mm. Microwave is a high-frequency electromagnetic wave that penetrates into the interior of a mineral to cause orientation polarization and deformation polarization of matter molecules. As the electrode changes, the direction of polarization also changes constantly, resulting in the self-heating effect of the mineral body, and the temperature rises. However, due to the different mineral properties of the ore, the absorbing properties are also different, resulting in ore. The temperature difference between each mineral in the mineral is different, and the thermal expansion coefficient of each mineral is also different. As a result, thermal cracking and the like occur, which causes microcracks in the mineral system and expands the original microcracks, thereby facilitating subsequent pulverization operations. Although microwave heating treatment has the incomparable advantages of traditional heating methods, the current theoretical research on microwave grinding is not deep enough. In the near future, microwave will play a huge role in reducing the energy consumption and steel consumption of grinding operations.
The so-called selective grinding is the use of selective dissociation of minerals and selective grinding of the grinding, the purpose is to cause some selectivity in grinding operations. The main purpose of the grinding operation is not to reduce the ore particle size, but to dissociate the useful minerals from the gangue minerals. The ultimate goal of grinding is to obtain the highest monomer dissociation with minimal energy input. Selective grinding is widely used in mining production such as metal ore, non-metallic minerals and coal mines, especially in the production practice of bauxite.
Because the ore materials of different particle sizes have different requirements on the grinding form, the coarse-grained materials are suitable for the grinding method based on impact crushing, and the fine materials should be ground by grinding. Micro-stage grinding is to install barrel linings with different surface shapes along the axial direction of the barrel of the ball mill. The surface of the ball mill is installed with a non-smooth lining to form a high steel ball drop height, resulting in impact pulverization. . A smoother cylinder liner is installed at the discharge end of the ball mill to form a lower steel ball drop height, resulting in grinding and pulverizing.
From the feeding end to the discharge end, the pulverized form gradually changes from impact pulverization to grinding pulverization and pulverization, so that the grinding form changes along the axis direction of the ball mill, and stage grinding is realized in a billiard mill. This can better meet the different needs of the ore material in different stages of the grinding process, different particle size composition, and meet the ore crushing law, thus improving the grinding efficiency. The implementation of the micro-stage grinding technology only requires the modification of the surface of the partial cylinder liner, which is simple and easy.
The cost of conventional grinding technology is quite high. Even if the grinding particle size can make useful minerals dissociate, the over-grinding phenomenon will occur, and many useful minerals will be lost in the slime. After several years of continuous development, the ultra-fine grinding technology of the mill has become one of the important deep processing technologies of industrial minerals and raw materials, which is of great significance to the development of modern high-tech industries.
Ultra-fine grinding technology is applied in the pretreatment of refractory gold ore. The gold is coated with pyrite. The gold ore contained in microscopic gold, sub-microscopic gold or solid solution is a kind of gold that is extremely difficult to dissolve and extract gold. ore. The key to gold extraction is to destroy the pyrite package and expose the gold to dissociation. The pyrite is very stable and difficult to decompose. With the development of ultra-fine grinding technology, it is possible to use an ultra-fine grinding to open the package of sulfide to dissociate the gold.
There are many types of internal stress of the ball mill, such as impact, extrusion, shearing, grinding, etc. Some of the stress electric energy consumption is large and the pulverizing efficiency is not high. Studies have shown that when the compressive stress with high pulverization efficiency is selected as the main stress, since the pressure pulverization process conforms to the smashing law of the material layer, when the pressure is small, the free loose material is first sufficiently dense, and when the pressure is increased, the squeezing particles are mutually The stress is transmitted, and when the strength value of the particles is exceeded, the mineral particles are broken and a large number of microcracks are generated. Adjusting and strengthening the energy input for different material characteristics, and also restraining the action area of the stress, so that the material passes through the stress zone regularly, and the mechanical energy is effectively converted into the pulverization energy, so that the particle becomes a blank product with low porosity. Through the subsequent process, a product with qualified particle size is obtained, thereby achieving the goal of high production and energy saving.
The high-pressure roller mill developed based on this theory has been applied to large-scale industrial production and has achieved good economic benefits. It can significantly improve the processing capacity of equipment systems, reduce the power consumption per unit, and save infrastructure investment and simplify the process. , reduce the number of broken sections, general ore materials can be used, the feed water content can reach 15%.
In the past 10 years, the new grinding equipment has been continuously introduced, with the aim of obtaining a larger crushing ratio and obtaining more fine-grained crushing products to reduce the particle size of the grinding material, save energy and reduce consumption, and at the same time carry out structural innovation, adopting new technologies, new materials improve traditional equipment to improve reliability, durability, performance and efficiency.
The development of the gyratory crusher has a history of 100 years. Due to its large processing capacity, large ore size, and ability to handle hard ore, it is still important equipment for crushing various hard materials in large mines and other industrial sectors. The rotary crusher has large production capacity, low unit power consumption and stable operation. It is suitable for processing sheet materials. The crushed product has a relatively uniform particle size and can be widely used for coarse crushing and medium crushing of various hardness ores. However, compared with the jaw crusher, the structure is complicated, the price is high, the maintenance is difficult, the repair cost is high, and the capital construction cost is high.
With the adoption of large-scale carrying equipment, the crushing mills ore feeding size has reached 1.2 to 2 m, which has promoted the development of the jaw crusher to large-scale. The compound pendulum crusher has the advantages of high efficiency and low price, occupies a large market share of the jaw crusher. With the promotion of conservation, energy-saving and efficient production methods, several new jaw crushers have also been successfully developed.
The spring cone crusher has been in existence for a hundred years. It was invented by the American Symons brothers using the principle of a gyratory crusher. So far, its structure has not changed much, its performance is stable, and it has a certain market share. In order to meet todays high throughput production, achieve high energy, achieve higher crushing ratio and finer product granularity, the new cone crusher has also been continuously developed and applied to production practice. For example, a hydraulic cone crusher that uses hydraulic instead of a spring and an inertia cone crusher that can replace the rough grinding operation have achieved good economic benefits in production and operation.
The high-pressure roller mill, also known as the roller crusher, works on the principle of material layer pulverization. It is a new type of high-efficiency energy-saving grinding equipment, which is gradually being applied and popularized at home and abroad. When the high-pressure roller mill was originally designed and applied, it was mainly used for the crushing of limestone and brittle metal ore with less hardness, and it was used for the middle and fine crushing of crushing operations. After years of promotion and development, it has been used in fine crushing of medium hardness and above, especially in the case of iron ore crushing, its technology has become increasingly mature, it has a large crushing ratio, fine product size, high efficiency, low energy consumption, etc. The utility model can also be applied to replace a rough grinding operation, and the ore can be crushed by a high-pressure roller mill to obtain a product of 3 to 10 mm in size, and the magnetite can be greatly improved after pre-magnetic separation. The grade of the mine has the characteristics of water saving, electricity saving and production increase. At present, the high-pressure roller mill is developing in the direction of large-scale, the diameter of the roller and the roll surface are further increased, the grain size range of the grinding is larger, and the throughput is also increased. The production practice shows that the single-machine production capacity of the high-pressure roller mill can reach 1,500 to 2 000 t / h, and the energy consumption of the crushed metal ore is 1.2 to 2. 8 kWh / t. Under the same conditions, the unit energy consumption is broken than the conventional one. The machine is 20% to 50% lower, the roll surface wear resistance is good, the service life of the inlaid hard alloy grain nail roll surface can reach 8 500 h, and the automation level is high. With the improvement of the performance of the high pressure roll mill, it is necessary for the metal mine. There will be broad application prospects.
The mill is further developed to a large scale. The change of the diameter of the mill has obvious changes for the grinding process. The large mill usually has a high specific crushing rate and can handle coarser grade materials. However, if the diameter of the mill is too large, the dead zone of the ball will increase. When the mill increases the processing capacity, it will also reduce the residence time of the mineral material, hindering the transfer of energy from the ball medium to the ore particles, resulting in a unit volume yield. The decline, the unit energy consumption of grinding products increased, so the development direction of the mill has been developed from large-scale to high-efficiency energy-saving.
Since the use of self-grinding and semi-self-grinding technology in the 1950s, it has grown into a mature, reliable and continuously applied technology. In the self-grinding process, the ore larger than 100 mm in the mill acts as a grinding medium. The ore material with less than 80 mm and more than 20 mm has poor grinding ability, and it is not easily broken by large ore materials, sometimes When the material is crushed, a steel ball of about 4% to 8% of the volume of the mill is often added to the mill, which improves the grinding efficiency of the mill, and thus semi-self-grinding occurs. The semi-autogenous mill belongs to a cylindrical mill with heavy load, low speed and large starting torque. Nowadays, both the new expansion and the renovation of the old factory, almost all use self-grinding, semi-self-grinding technology, self-grinding, The semi-self-grinding technology eliminates the two-stage crusher and the screening equipment, simplifies the process and improves the operating conditions, which not only reduces the capital cost of the construction, but also reduces the production and operation costs, and also facilitates automation.
The ball mill is a traditional material crushing device. It has a history of more than 100 years. It is still important equipment for fine powdering of solid materials. It is widely used in metallurgy, chemical industry, cement, ceramics, construction, electricity and In the industrial sectors such as national defense, dry and wet grinding of various ores and materials is possible. In recent years, the development of ball mills has focused on energy saving and consumption reduction, continuously improving and perfecting the grinding machine transmission mode, researching and developing new lining plates and grinding media, striving to achieve automatic control of the grinding process, and improving on the premise of ensuring grinding grain size. The processing capacity and grinding efficiency of the mill.
As we all know, the lining of the ball mill is a key part of the mill that can achieve high efficiency, energy saving and consumption reduction. After research and development, it has made good progress.
-The angle spiral lining is also called energy-saving lining. After using this lining, the unit output power consumption is reduced by 10% to 25%, the mill output is increased by 15% to 20%, and the unit output ball consumption is reduced by 10% to 20%. It has the advantages of stable operation, less product pulverization, less noise, etc., and is especially suitable for crushing operations in cement production;
-The rubber lining is a corrosion-resistant and wear-resistant non-metallic material lining. Compared with the manganese steel lining, it has the advantages of a lightweight, low energy consumption, high output and low noise.
-On the basis of the rubber lining, a composite magnetic lining has been developed. This lining is magnetically adsorbed on the surface of the lining to adsorb a layer of magnetic particles and dielectric fragments to form a protective layer to extend the service life of the lining. It is almost half lighter than ordinary manganese steel lining, and can be directly adsorbed on the inner surface of the mill barrel without bolt fixing, which greatly reduces the workload of installation and maintenance, not only reduces energy consumption, but also increases the processing capacity of the mill.
The rod mill is developed on the basis of the ball mill. It has the advantages of reliable processing technology, low investment, less auxiliary equipment and simple process flow. It can be combined with the ball mill to form a different grinding process. The rod mill mainly grinds the ore by the pressure and the grinding force of the grinding rod. When the rod hits the ore, it first hits the coarser grade ore, and then pulverizes the smaller-sized material, between the rod and the rod. When the rod is in contact with the wall, the coarser-grained ore particles are mixed with it, which acts as a bar sieve. The finer-grained material can pass through the gap between the rod and the rod, which is beneficial to the clamp. The coarser-grained material also allows the coarser-grained ore particles to be concentrated in the place where the grinding media strikes. Therefore, the rod mill has the function of selective grinding, and the product has a uniform particle size and less pulverization.
The vertical spiral mixing crusher is a new type of high-efficiency energy-saving grinding equipment successfully developed by Changsha Research Institute of Mining and Metallurgy. Its grinding effect is mainly grinding and stripping, as well as a small amount of impact and shearing, so that the original material can be kept. Lattice shape, make full use of energy to effectively grind the material, because in the fine grinding and ultra-fine grinding, friction grinding is the most effective grinding method, and has been used in the regrind or fine grinding operation of metal mines.
High-speed impact pulverizer refers to an impact pulverizing device that strongly impacts the material around a horizontal or vertical high-speed rotating body (rotor, hammer, blade), which can crush materials below 8 mm to 10 m at 70%. the above. The device can be applied to ultra-fine pulverization of non-metal such as talc, clay, barite, calcium carbonate, mica, and graphite.
There are many factors affecting the grinding efficiency, including the nature of grinding feed, the size of the ore, the filling rate of the steel ball, the size of the steel ball and the ratio, the ball filling system, the grinding system, the grinding process, the mill operation, and the classification. Factors such as efficiency and amount of sand return, but these factors are not independent of each other and have a certain impact on each other.
The mechanical properties of the ore, such as hardness, toughness, dissociation and structural defects, determine the grindability of the ore, which determines the difficulty of grinding. The small grinding degree indicates that the ore is easy to grind, the smaller the wear of the ore on the mill, the lining and the grinding medium, the smaller the power consumption is consumed; on the contrary, if the grinding degree is large, the wear of the mill And power consumption is big. The nature of the ore will directly affect productivity and the impact on grinding operations is of paramount importance. In the modern grinding operation, a grinding aid process has been added to add some specific chemicals to the grinding process to reduce the grindability of the ore and increase the productivity of the mill.
The grain size of the mill has a great influence on the grinding efficiency of the mill. Generally speaking, the smaller the grain size, the smaller the work done by the mill on the ore; on the contrary, the larger the grain size, the mill The work done on the ore is greater. The crushing of ore by steel ball is a random crushing, and the crushing efficiency is very low. Some researches have pointed out that the crushing efficiency of the ball mill is only 6% to 9%. It can be seen that the grinding grain size has a great influence on the mill. In order to achieve the final grinding fineness, it will inevitably increase the workload of the ball mill, and the energy consumption and power consumption of the ball mill will also increase.
There is a close relationship between the rotation rate and the filling rate of the mill. The two are related to each other. Generally speaking, once the mill is installed, its rotation rate is fixed, it will not change easily, and the operation of changing the rotation rate is compared. It is cumbersome, so in actual production, the transfer rate is generally not analyzed as a factor affecting the grinding efficiency. It is only necessary to analyze the suitable ball filling rate at a certain speed. When the transfer rate is constant, the filling rate is large, the steel ball hits the material more frequently, the grinding area is large, the grinding effect is strong, but the power consumption also increases, and the filling rate is too high, which also affects the steel. The movement state of the ball reduces the impact on the large material; on the contrary, if the filling rate is small, the grinding area is small, the grinding effect is relatively weak, but the power consumption and energy consumption are also small. Therefore, at the production site, whether the filling rate is appropriate has a great influence on the grinding efficiency of the plant.
In the mill, the steel ball and the mill are in point contact. When the ball diameter is too large, the crushing force is also large, so that the ore is broken along the direction of the penetrating force, instead of being broken along the crystal plane of different minerals with weak bonding force, resulting in no fracture. Selective. In the case of the same filling rate of the steel ball, the ball diameter is too large, resulting in too few steel balls, low breaking probability, severe crushing, and uneven product size; on the contrary, if the steel ball is too small, its crushing effect on ore is small. The grinding efficiency is low, so the precise steel ball size and ball ratio have a great influence on the grinding efficiency.
The main function of the ball mill liner is to protect the mill. When the mill is running, the steel balls and materials inside the mill are thrown or slid by the liner to a certain height, and the material is ground and pulverized. It will also be affected by the impact, sliding and rolling of steel balls and materials, and will also be affected by temperature. Therefore, the main form of wear of the lining plate is abrasive wear under a small number of times of energy, so which material lining is selected, Reducing its wear and tear is always an important issue for ball mills. At present, there are three main types of lining materials widely used: high manganese steel; alloy wear-resistant steel; high chromium cast iron.
High manganese steel has good wear resistance and good economic applicability, but low yield strength, suitable for medium and Use under high impact load wear conditions. -Alloy wear-resistant steel has a higher comprehensive performance than high-manganese steel and is suitable for medium impact wear conditions. -High-chromium cast iron has a higher wear resistance than the former two and is more widely used.
Grinding concentration is also an important factor affecting the grinding efficiency. Its size will affect the specific gravity of the slurry, the adhesion of the ore around the steel ball and the fluidity of the slurry. When the grinding concentration is low, the fluidity of the slurry is fast, and the adhesion of the material around the steel ball is low, so that the impact and grinding effect of the steel ball on the material are weakened, and the grinding efficiency is low. When the grinding concentration is high, the adhesion of the material around the steel ball is good, the impact and grinding effect of the steel ball on the material is better, but the slurry fluidity is poor, the over-grinding is more serious, and it is not conducive to improving the processing capacity of the mill. Therefore, determining the optimum grinding concentration will have an important impact on the grinding efficiency.
For a long time, people tend to pay attention to the realization of grinding purposes, while ignoring the means and methods of grinding, patronizing the grinding grain size of the pursuit of requirements, and neglecting the monomer solution of various useful materials of ore containing various metals. The difference in the degree of separation will cause some minerals to be pulverized and some minerals to be insufficiently pulverized. In this case, if the conventional rough grinding process is still used, the grinding and sorting effects will not be good.
The classifier and the grinding machine work in a closed circuit, which can control the grain size of the grinding product and increase the productivity of the mill. Therefore, the classification efficiency has a certain influence on the grinding efficiency. When the classification efficiency is high, the qualified grain grade products can be eliminated in time. Avoid over-pulverization and reduce energy consumption; when the classification efficiency is low, the products that reach the qualified size can not be discharged in time and returned to the mill for re-grinding, which can easily cause over-grinding and affect the later selection effect.
The return-sand ratio is the ratio of the amount of sand returned by the ball mill to the ore and ore. The effect of graded anti-sand is not only to return the unqualified coarse particles, but also another important role to make the ball mills ore thickening and let the steel ball High efficiency crushing over the entire axial length of the mill increases the productivity of the mill. Under normal circumstances, the amount of sand return should not exceed 500%, and the second section should not exceed 690%.
There are many variables in the operation of the grinding classifier, and the change of one factor can cause successive changes of many factors. The manual operation can not keep up with this change, can not meet the requirements of the production process, and adopt automatic control to make The grinding grade is maintained in a stable and suitable state, thereby increasing productivity and reducing energy consumption.
Chinas grinding equipment has undergone considerable development through technology introduction, technical cooperation, digestion and absorption, and self-development. At present, China has a wide variety of grinding equipment, complete varieties, continuous improvement in manufacturing quality, and increasing production year by year. It has become one of the countries with the most productive grinding equipment in the world.
The process flowsheet of Uranium generally outlines the latest proven processes for uranium concentration known as Resin In Pulp more commonly referred to as RIP To date it is not applicable to ores containing vanadium, where the vanadium must be recovered.
Ore is delivered to the receiving hopper, which is provided with a grizzly. Feed to the primary Jaw Crusher is metered by a Apron Ore Feeder which discharges over a grizzly. The oversize is crushed and conveyed with the undersize from the grizzly to a Dillon Vibrating Screen. The screen scalps out the fines ahead of the secondary cone crusher.
A permanent magnet protects the secondary crusher from tramp iron. Closed circuit is preferred, as it minimizes the size of the primary sampler and possibly one stage of reduction in the sampling sections. Smaller tonnage mills can usually eliminate the secondary stage of crushing and feed material as coarse as 2 to the rod or ball mill. This also depends on hardness of the ore.
Size reduction with Crushing Rolls or Jaw Crushers, and controlled feeding takes place between each stage of sampling. The secondary and tertiary Vezin Samplers are fitted with four cutters to enable the operator to obtain any quantity of sample desired, regardless of the size of shipment. The Coffee Mill makes the final size reduction.
Ore is fed at a controlled rate by Adjustable Stroke or Variable Speed Ore Feeders. A weightometer records tonnage and a Automatic Sampler provides systematic sampling. Normal procedure in the RIP Process is to grind ores to 28 or 35 mesh in a Rod Mill or Ball Mill, which is operated in closed circuit with a Rake Classifier. Classifier overflow must be in excess of 50% solids due to subsequent acid leaching.
Acid leaching of high lime ores is extremely costly. Ores of this type are sent to Sub-A Flotation Machines for removal of a selective lime concentrate. The lime flotation concentrate would pass to a Thickener for dewatering prior to carbonate leaching to recover uranium. Underflow from the thickener would be metered by a Adjustable Stroke Diaphragm Pump. Flotation carried on at high density in Cells would eliminate subsequent thickening of the flotation tailings prior to leaching operations.
Pulp at approximately 55% solids is pumped to the Acid Leaching Section. Sufficient concentrated H2SO4 is added to bring the pH below 1.0. Usually acid is stage added over the first five hours leaching time. Most ores also require the addition of an oxidizing agent to aid in the dissolution of uranium. Normal leaching period is 11 hours, however, most plants design for 15 hours.
Heavy duty Acid Proof Agitators are used in the leaching sections. Rubber or neoprene covered shafts and propellers are furnished with acme threads. Worn propellers are replaced by merely unscrewing from the shaft. This minimizes delay and eliminates having to send the shaft and propeller to the rubber company for covering, as a unit. Tanks are either wood, rubber or neoprene lined.
In order to minimize wear on this relatively coarse feed, the units are equipped with Turbine Type Propellers. The diameter approaches one-third the diameter of the tank which allows for slow speed rotation and thereby maximum wearing life. Powdered iron is sometimes added by means of a Cone Type Feeder, to the final agitator discharge, to reduce ferric iron to ferrous, to prevent iron precipitation in the R.I.P. section.
The agitator discharge flows to the number one Acid ProofRake Classifier, usually rubber or neoprene covered tank and mechanism with stainless steel 3/16 rake blades. A 325-mesh overflow is desired. The five classifier units are operated in series according to standard counter current practice with new water being added to the last unit. The sand fraction discharges to a Repulping Conditioner, where it is mixed with the slime tailings fraction, and a SRL Rubber Lined Pump delivers to the tailings area.
The number one classifier overflow is pumped through a cyclone to further remove any +325 mesh solids which might have overflowed the classifier. The coarse fraction is sent to Rake Classifier No. 2, so as not to overload the No. 1 unit. The slime fraction from the cyclone flows to a Dillon Vibrating Screen as a safety factor for removal of wood chips, paper, etc.
The pregnant slime is introduced to this circuit for the recovery of uranium values by absorption. Size of cells runs from 4 to 6 square and number of cells to bank from 2 to 4. There are usually 11 to 14 banks per section.
The R.I.P. Cells consist of trough type tanks in which screen covered (20 mesh) baskets are moved slowly up and down. The baskets contain resin in the form of beads. The pregnant slime flows through the tank and the uranium is absorbed in the resin. When the bank becomes fully loaded it is removed from the absorption cycle.
A nitrate stripping solution is then passed through the cells and the uranium removed in solution from the beads. This pregnant solution is referred to as the eluate. A water wash is used to remove all traces of nitrate, and the bank of cells is then ready to be placed back on the absorption cycle. The nitrate wash water goes to stripping solution makeup tanks.
Vertical Sand Pumps, (not shown on flowsheet) are used to handle flows between banks so that as cycles change, flows may be directed to any bank desired. The problem calls for and this pump is designed to handle widely varying flows without surging, and to operate dry without damage to the pump.
The eluate is then pumped to sand filters or other methods of classification to remove all solids prior to uranium precipitation. Uranium precipitation is usually accomplished on a batch basis by the addition of MgO to the solution in light duty Agitators. The precipitate is then pumped to a Thickener for removal of solution. The thickener overflow is pumped through a filter press to insure complete recovery of uranium precipitate. The filtrate is reused at this point as makeup for the stripping solution.
Thickener underflow is controlled with a Adjustable Stroke Diaphragm Pump which discharges to a Washing Drum Filter. The filter cake is repulped, filtered and dried. In large plants continuous steam heated Drum Dryers are used.
During the past few years, the solvent extraction process has been applied commercially to the recovery of uranium. Considerable research is now underway to adapt this process to the recovery of other minerals. The process provides a simple, selective and inexpensive method of upgrading uranium solutions ahead of the precipitation sections.
To accomplish the process objectives, a counter- current decantation circuit is used. The pregnant aqueous solution flows in one direction and eventually leaves the unit as barren aqueous or raffinate. The organic solvent flows in the opposite direction and eventually leaves the unit as pregnant organic which is pumped to the stripping section for further processing. The organic solvent is a mixture of organic materials and acarrier which is usually high flash point kerosene. The solvent extraction process involves the proper mixing of the uranium (or other mineral) pregnant aqueous liquor with the organic solvent in order totransfer the uranium values from a large volume of aqueous liquor to a much smaller volume of organic solvent. When the organic solvent is properly mixed with the pregnant aqueous it picks up the uranium from the aqueous as uranium has a greater affinity for the organic than it has for the aqueous.
The flow from the mixer process passes to the settling process where the lighter kerosene containing the organic solvent and uranium rises to the top and the heavier aqueous settles to the bottom. Usually five stages of mixing and settling are used to effect maximum extraction of uranium and a compact mixer-settler unit has many cost-reducing advantages.
In 1956, Kerr McGee Oil Industries, one of the pioneers in the development and the commercial application of the solvent extraction process in uranium, designed a continuous solvent extraction circuit similar to the one illustrated in this design study. It accomplished the following:
All of the equipment used in the solvent extraction process for uranium must be acid proof to withstand the corrosive nature of the 1.0 pH pregnant aqueous solution used in the process. The presence of kerosene used as the solvent carrier eliminates the possible use of rubber or neoprene in the design of equipment. Normally, for mechanisms and hardware, stainless steel 316 has been found satisfactory, except where chlorides are present in the solutions. With the presence of chlorides, polyvinyl chloride has been found to be satisfactory. Wood tanks with wood mixer and settler partitions have worked very well andinvolve comparatively little expense.
Since the Kerr McGee installation, similarly designed systems have been installed at the Gunnison Mining Company, Fremont Minerals, Inc., and Mines Development, Inc., uranium concentrators. Minor modifications have been made in the various installations. However, probably the most significant change in the basic design was the method for the organic recirculation as illustrated in this Design Study.
Flow of the aqueous through the system is by gravity. The organic is advanced through the circuit by means of an airlift or pumps. It is necessary to have a high ratio of the organic solvent to the aqueous within the mixing chamber for proper operation. The organic is recirculated from the settler compartment back to the mixer compartment. This is done by means of an inexpensive Pumping Turbine installed on the bottom of the Solvent Extraction Mixer shaft. An overflow weir is installed in the settler section to allow a portion of the organic in the settler to flow to the bottom of the mixer compartment where it is lifted by the pumping turbine and recirculated in the mixing compartment.
Gravity flow of the aqueous through the unit is controlled by weirs between each settler and mixer. There is a 6 differential in elevation between the weirs in each preceding compartment. The flow of the lighter organic passes to the boot of an airlift where it is advanced to the next mixer compartment.
The interface level between the organic and aqueous phases in the settlers is controlled automatically by the position of the weirs between the settlers and mixers. A monometer effect, using the difference in gravity between the lighter weight organic and the heavier aqueous phases, prevents the organic from passing out to the tailings as it would never have sufficient weight to force the column of aqueous over the weir. On the same basis, should the aqueous level start to rise in the settler it causes a higher flow of aqueous over the weir and thereby maintains the proper level.
Although basic porphyry copper flotation and metallurgy has remained virtually the same for many years, the processing equipment as well as design of the mills has continually been improved to increase production while reducing operating and maintenance costs. Also, considerable attention is paid to automatic sensing devices and automatic controls in order to assure maximum metallurgy and production at all times. For simplicity in this study most of these controls are not shown.Many of the porphyry copper deposits contain molybdenite and some also contain lead and zinc minerals.
Even though these minerals occur in relatively small amounts they can often be economically recovered as by-products for the expense of mining, crushing, and grinding is absorbed in recovery of the copper.
Because the copper in this type of ore usually assays only plus or minus 1% copper, the porphyry copper operations must be relatively large in order to be commercial. The flowsheet in this study illustrates a typical 3,000 ton per day operation. In general most operations of this type have two or more parallel grinding and flotation circuits. For additional capacity, additional parallel circuits are installed.
The crushing section consists of two or three crushing stages with the second or third stages in either closed or open circuit with vibrating screens. Generally, size of the primary crusher is not determined by capacity but by the basic size of the mine run rock. The mine-run ore is normally relatively large as most of the porphyry mines are open pit.The crushing section illustrated is designed to handle the full tonnage in approximately 8 to 16 hours thus having reserve capacity in case of expansion.
Many mills store not only the coarse ore but also the fine ore in open stockpiles using ore as the side walls and drawing the live ore from the center. During prolonged periods of crusher maintenance the ore walls can be bulldozed over the ore feeders to provide an uninterrupted supply of ore for milling.
As it is shown in this study the or 1 crushed ore is fed to a rod mill operating in open circuit and discharging a product approximately minus 14-mesh. The discharge from this primary rod mill is equally distributed to two ball mills which are in closed circuit with SRL Rubber Lined Pumps and two or more cyclone classifiers. The rod mill and two ball mills are approximately the same size for simplified maintenance.
Porphyry copper ores, usually medium to medium hard, require grinding to about 65-mesh to economically liberate the copper minerals from the gangue. Although a clean rougher tailing can often be achieved at 65-mesh the copper mineral is not liberated sufficiently to make a high grade copper concentrate, thus some form of regrinding is necessary on the rougher flotation copper concentrate. It is not unusual to grind the rougher flotation concentrate to minus 200-mesh for more complete liberation of mineral from the gangue.
The cyclone overflow from each ball mill goes to a Pulp Distributor which distributes the pulp to two or more parallel banks of Flotation Cells. These distributors are designed so that one or more flotation banks can be shut down for maintenance or inspection and still maintain equal distribution of feed to the remaining banks.
In some cases it is beneficial to have conditioning before flotation, but this varies from one operation to another and it is not shown in this flowsheet. Ten or more Free-Flow Flotation Cells are used per bank and these cells are divided into groups of four or six cells with an intermediate step-down weir between groups. Free-Flow Flotation Cells are specified, as metallurgy is extremely good while both maintenance and operating expenses are traditionally low. One or more Free-Flow Mechanisms can be stopped for inspection or even replaced for maintenance without shutting down the bank of cells.
The concentrates from rougher flotation cells are sent directly to regrind. Often the grind is 200-mesh. After regrind is flotation cleaning. In some cases the concentrate from the first three or four rougher flotation cells can be sent directly to cleaning without regrinding.
After the rougher flotation concentrate is reground it is cleaned twice in additional Free-Flow Flotation Machines with the recleaned concentrate going to final concentrate filtration or, as the metallurgy dictates, to a copper-moly separation circuit.
The thickening and filtering is similar to other milling operations, however, as the porphyry copper installations are often in arid areas, the mill tailing is usually sent to a large thickener for water reclamation and solids go to the tailings dam.
Automatic controls are usually provided throughout modern plants to measure and control pulp flow, pH and density at various points in the circuit. Feed and density controls are relatively common and the newer installations are using automatic pulp level controls on flotation machines and pump sumps. Automation is also being applied to the crushing systems.
The use of continuous on stream X-ray analysis for almost instantaneous metallurgical results is not shown in thus study but warrants careful study for both new and existing mills. Automatic sampling of all principal pulp flows are essential for reliable control.
The flowsheet in this study illustrates the modern approach to porphyry copper treatment throughout the industry. Each plant will through necessity have somewhat different arrangements or methods for accomplishing the same thing and reliable ore test data are used in most every case to plan the flowsheet and design the mill.
In most plants engaged in the flotation of ores containing copper-bearing sulphide minerals with or without pyrite, pine oil is employed as a frother with one of the xanthates or aerofloat reagents or a combination of two or more of them as the promoter. Lime is nearly always used for maintaining the alkalinity of the circuit and depressing any pyrite present. The reagent consumption is normally within the following limits
While good results are often obtained with ethyl xanthate alone as a promoter, the addition of a small quantity of one of the higher xanthates is frequently found to improve the recovery of those minerals that are not readily floated by the lower xanthate, especially those that are tarnished or oxidized, but since the action of a higher xanthate is, as a rule, more powerful than that of the ethyl compound, it is usually best to add no more of the former reagent than is necessary to bring up the less readily floatable minerals, controlling flotation with the less powerful and more selective lower xanthate. Better results are obtained with some ores by replacing the higher xanthate with one of the dithiophosphates, flotation being controlled, as before, with ethyl xanthate. Sometimes a dithiophosphate can be effectively used without the xanthate, although the dual promotion method is more common. A rule of thumb system for the selection of these reagents cannot be laid down as the character of the minerals differs so widely in different ores ; the best combination can only be found by experiment.When aerofloat is employed alone as the promoter, the reagent mixture is somewhat different from that given above. A reliable average consumption is difficult to determine as the plants working on these lines are few in number, but the following is what would normally be expected.If this combination of reagents gives results equal to those obtainable with a xanthate mixture, its employment has these advantages over the latter method: The control of flotation is not so delicate as with xanthates, it has less tendency to bring up pyrite, and, if selectivity is not required, the circuit may be neutral or only slightly alkaline.
When the ore is free from pyrite, the function of the lime, whatever the reagent mixture, is to precipitate dissolved salts and to maintain the alkalinity of the pulp at the value which has been found to givethe best results ; soda ash is seldom employed for this purpose. When pyrite is present, lime performs the additional function of a depressor, the amount used being balanced against that of the promoterthat is, no more lime should be added than is required to prevent the bulk of the pyrite from floating, as any excess tends to depress the copper minerals, and no more of the promoter should be employed than is needed to give a profitable recovery of the valuable minerals in a concentrate of the desired grade, since any excess tends to bring up pyrite. In many cases a more effective method of depressing pyrite is to add a small quantity of sodium cyanidee.g., 0.05-0.10 lb. per tonin conjunction with lime, less of the latter reagent then being necessary than if it were used alone.
It is not often that a conditioning tank has to be installed ahead of the flotation section in the treatment of sulphide copper ores, as the grinding circuit usually provides suitable points for the introduction of the reagents. The normal practice is to put lime into the primary ball mills and to add xanthates at the last possible moment before flotation, while aerofloat and di-thio-phosphates are preferably introduced at some point in the grinding circuit, since they generally need an appreciable time of contact as compared with xanthates. There is no special place for the addition of pine oil, but care should be taken if it is put into the primary ball mills, as a slight excess may cause an undue amount of froth to form in the classifiers.
In a plant where the primary slime is by-passed round the grinding circuit, it is necessary to ensure that this portion of the pulp receives its correct proportion of and contact time with the reagents.
As regards flotation installations, the present tendency is to employ machines of the air-lift or Callow-Maclntosh rather than of the subaeration type. While two stages of cleaning (circuits 10 and 11) are sometimes essential to the production of a clean final concentrate, circuits 8 and 9 comprising a single stage of cleaning are probably the most widely used. Occasionally the primary machines can be run as rougher-cleaner cells (circuit No. 5), particularly when they are of the air-lift or subaeration type. This method, however, is not often employed, although its use is more common in the flotation of copper sulphide minerals than of any other class of ore ; a stage of cleaning is preferable as providing greater lattitude of control.
Two variations of normal procedure are worth notice. In one or two plants employing two-stage grinding, improved results have been obtained by separating the slime from the primary ball mill circuit and sending it direct to a special flotation section. This method is useful when the feed to the flotation plant contains an appreciable quantity of fines, which, due generally to oxidation through exposure, require different treatment from the unweathered part of the ore. Such fines are usuallyfriable and can be separated as slime from the primary grinding circuit without the inclusion of an undue proportion of unoxidized material, the bulk of which thus passes to the secondary grinding circuit and thence to its own division of the flotation plant.
The second variation consists of grinding the rougher concentrate before cleaning. The method is applicable to an ore in which the copper- bearing minerals are so intimately associated with pyrite that very fine grinding is necessary to liberate them completely. It is often possible, after grinding such an ore to a comparatively coarse mesh, to make a profitable recovery of the copper in a low-grade concentrate which does not represent too large a proportion, say 30% or less, of the total weightof the feed. The concentrate can then be reground and refloated with the production of a high-grade copper concentrate together with a low- grade pyritic tailing suitable for return to the roughing circuit. This method is likely to be less costly than one involving the fine grinding of the whole ore. No standard system can be given for handling the various products as their disposal depends so much on the occurrence of the minerals and the efficiency of the regrinding operations, but a typical flow sheet is illustrated in circuit No. 12 (Fig. 60). It is diagrammatic to the extent that the thickener and regrinding unit may receive its feed from several roughing machines and deliver its discharge to a number of cleaning cells. It is usual to dewater the rougher concentrate and return the water to the primary circuit for two reasons : First, to supply the regrinding mill with a thick enough pulp for efficient operation, and, secondly, as far as possible to prevent the reagents used in the roughing circuit from entering the cleaning section.
In normal practice a recovery of over 90% of the copper which is present as a sulphide is generally possible, whatever the flotation process or circuit employed. As regards the average grade of concentrate, no more can be said than that it depends on the class of the copper-bearing minerals present and their mode of occurrence and on the character of the gangue. It usually contains over 20% of copper, but a difficult chalcopyritic ore may yield a concentrate with less than that percentage, while it is theoretically possible to obtain one running over 75% should the mineral consist entirely of pure chalcocite.
The flotation of native copper ores is nearly always preceded by gravity concentration in jigs and tables not only because the combined process is more economical as regards costs, but also because the copper often occurs as large grains which flatten out during grinding and cannot be broken to a size small enough for flotation. The flow sheet depends on the mode of occurrence of the mineral. The tailings from some of the gravity concentration machines may be low enough in value to be discarded, but those products which still contain too much copper to be sent to waste are thickened and reground, should either operation be necessary, and then floated with pine oil and a xanthate or aerofloat reagent in a neutral or slightly alkaline circuit. The reagent consumption is approximately the same as that given for the treatment of copper- bearing sulphides. While a pine oil, lime, and ethyl xanthate mixture has proved satisfactory, better results have sometimes been obtained by the substitution of aerofloat and sodium di-ethyl-di-thio-phosphate, soda ash being used instead of lime on account of its gangue deflocculating properties. On the average 0-12 lb. per ton of aerofloat and 0.03 lb. of the di-thio-phosphate are substituted for 0.1 lb. of xanthate.
Since a high-grade concentrate is desired in order to keep smelting costs as low as possible, the circuit usually comprises two stages of cleaning. In most plants flotation is carried out in mechanically agitated machines.
The problem of the flotation of oxidized copper ores has not yet been solved. One or two special processes are in operation for the flotation of malachite and azurite, but none of them has more than a limited application; nor has any method been worked out on a large scale for the bulk flotation of mixed oxidized and sulphide copper minerals when the former are present in the ore in appreciable quantity.
For thousands of years the word gold has connoted something of beauty or value. These images are derived from two properties of gold, its colour and its chemical stability. The colour of gold is due to the electronic structure of the gold atom, which absorbs electromagnetic radiation with wavelengths less than 5600 angstroms but reflects wavelengths greater than 5600 angstromsthe wavelength of yellow light. Golds chemical stability is based on the relative instability of the compounds that it forms with oxygen and watera characteristic that allows gold to be refined from less noble metals by oxidizing the other metals and then separating them from the molten gold as a dross. However, gold is readily dissolved in a number of solvents, including oxidizing solutions of hydrochloric acid and dilute solutions of sodium cyanide. Gold readily dissolves in these solvents because of the formation of complex ions that are very stable.
Gold (Au) melts at a temperature of 1,064 C (1,947 F). Its relatively high density (19.3 grams per cubic centimetre) has made it amenable to recovery by placer mining and gravity concentration techniques. With a face-centred cubic crystal structure, it is characterized by a softness or malleability that lends itself to being shaped into intricate structures without sophisticated metalworking equipment. This in turn has led to its application, from earliest times, to the fabrication of jewelry and decorative items.
The history of gold extends back at least 6,000 years, the earliest identifiable, realistically dated finds having been made in Egypt and Mesopotamia c. 4000 bc. The earliest major find was located on the Bulgarian shores of the Black Sea near the present city of Varna. By 3000 bc gold rings were used as a method of payment. Until the time of Christ, Egypt remained the centre of gold production. Gold was, however, also found in India, Ireland, Gaul, and the Iberian Peninsula. With the exception of coinage, virtually all uses of the metal were decorativee.g., for weapons, goblets, jewelry, and statuary.
Egyptian wall reliefs from 2300 bc show gold in various stages of refining and mechanical working. During these ancient times, gold was mined from alluvial placersthat is, particles of elemental gold found in river sands. The gold was concentrated by washing away the lighter river sands with water, leaving behind the dense gold particles, which could then be further concentrated by melting. By 2000 bc the process of purifying gold-silver alloys with salt to remove the silver was developed. The mining of alluvial deposits and, later, lode or vein deposits required crushing prior to gold extraction, and this consumed immense amounts of manpower. By ad 100, up to 40,000 slaves were employed in gold mining in Spain. The advent of Christianity somewhat tempered the demand for gold until about the 10th century. The technique of amalgamation, alloying with mercury to improve the recovery of gold, was discovered at about this time.
The colonization of South and Central America that began during the 16th century resulted in the mining and refining of gold in the New World before its transferal to Europe; however, the American mines were a greater source of silver than gold. During the early to mid-18th century, large gold deposits were discovered in Brazil and on the eastern slopes of the Ural Mountains in Russia. Major alluvial deposits were found in Siberia in 1840, and gold was discovered in California in 1848. The largest gold find in history is in the Witwatersrand of South Africa. Discovered in 1886, it produced 25 percent of the worlds gold by 1899 and 40 percent by 1985. The discovery of the Witwatersrand deposit coincided with the discovery of the cyanidation process, which made it possible to recover gold values that had escaped both gravity concentration and amalgamation. With E.B. Millers process of refining impure gold with chlorine gas (patented in Britain in 1867) and Emil Wohlwills electrorefining process (introduced in Hamburg, Ger., in 1878), it became possible routinely to achieve higher purities than had been allowed by fire refining.
The major ores of gold contain gold in its native form and are both exogenetic (formed at the Earths surface) and endogenetic (formed within the Earth). The best-known of the exogenetic ores is alluvial gold. Alluvial gold refers to gold found in riverbeds, streambeds, and floodplains. It is invariably elemental gold and usually made up of very fine particles. Alluvial gold deposits are formed through the weathering actions of wind, rain, and temperature change on rocks containing gold. They were the type most commonly mined in antiquity. Exogenetic gold can also exist as oxidized ore bodies that have formed under a process called secondary enrichment, in which other metallic elements and sulfides are gradually leached away, leaving behind gold and insoluble oxide minerals as surface deposits.
Endogenetic gold ores include vein and lode deposits of elemental gold in quartzite or mixtures of quartzite and various iron sulfide minerals, particularly pyrite (FeS2) and pyrrhotite (Fe1-xS). When present in sulfide ore bodies, the gold, although still elemental in form, is so finely disseminated that concentration by methods such as those applied to alluvial gold is impossible.
Native gold is the most common mineral of gold, accounting for about 80 percent of the metal in the Earths crust. It occasionally is found as nuggets as large as 12 millimetres (0.5 inch) in diameter, and on rare occasions nuggets of native gold weighing up to 50 kilograms are foundthe largest having weighed 92 kilograms. Native gold invariably contains about 0.1 to 4 percent silver. Electrum is a gold-silver alloy containing 20 to 45 percent silver. It varies from pale yellow to silver white in colour and is usually associated with silver sulfide mineral deposits.
Gold also forms minerals with the element tellurium; the most common of these are calaverite (AuTe2) and sylvanite (AuAgTe4). Other minerals of gold are sufficiently rare as to have little economic significance.
Of the worlds known mineral reserves of gold ore, 50 percent is found in South Africa, and most of the rest is divided among Russia, Canada, Australia, Brazil, and the United States. The largest single gold ore body in the world is in the Witwatersrand of South Africa.
Crushing of mica prior to pneumatic processing technique were studied for concentrating coarse mica. Muscovite and phlogophite are the two major commercial mica minerals. However, this research was conducted exclusively with muscovite, and throughout this paper mica will mean muscovite. Two primary forms of commercial mica are (1) sheet mica and (2) scrap and flake mica.
Based on the quantity of visible inclusions and structural imperfections, the American Society for Testing and Materials (ASTM) has designated 12 quality groups for sheet mica. These designations range from black- and red- stained to perfectly clear. ASTM has also designated 12 grades of sheet mica based on the size of the maximum usable rectangle: sizes range from grade 6 (with 1 usable square inch) to grade OOEE special (with 100 usable square inches).
Scrap and flake mica is generally any mica of a quality and size that is not suitable for use as sheet mica. Most scrap and flake mica is recovered from schists, pegmatites, and, occasionally, as a secondary product from the beneficiation of feldspar and kaolin. Such mica is processed into ground mica for various end uses. Coarse, dry-ground mica of 5-mesh size is used in oil-well drilling mud to overcome lost circulation of the drilling fluid. Decorative finishes on concrete, stone, and brick are made with 16-mesh mica.
Twenty- and thirty-mesh mica is used to prevent sticking and for weatherproofing in the manufacture of roll roofing and shingles. Wallboard joint cements contain 100- and 200-mesh mica to eliminate cracking and reduce shrinking. Very fine mica is used in paints to improve exterior durability.
The United States is almost totally dependent on imports for its sheet mica. The high cost of skilled labor needed to mine and beneficiate sheet mica is prohibitive for many U.S. companies. The domestic supply of scrap and flake mica is reported to be adequate, although there is a short supply of high-quality scrap and flake mica for mica paper production. The Bureau of Mines investigated the pneumatic processing method as an alternative means to recover both sheet mica and coarse flake mica.
Sheet mica is selectively mined and beneficiated by hand. Scrap and flake mica can be recovered by several methods. The simplest method separates the mica from its host rock by differential crushing and screening in washer plants. Crushing, which has little effect on the size of the mica because of its platy, flexible characteristics, can effectively recover plus -inch mica. A second method utilizes screens, classifiers, and Humphreys spirals to concentrate the mica from the ground ore, which permits the recovery of a finer size mica. Also, flotation methods can recover minus 20-mesh mica with recoveries ranging from 70 to 92 percent.
An alternative technique, designed by the Bureau of Mines, uses crushers, screens, and zigzag air classifiers to concentrate mica. Sheet or flake mica has two dimensions many times larger than the third dimension. After screening the ore into close size fractions, the mica sheets or flakes are significantly lighter than the gangue particles of the same size fraction.
Air classification separates the flat, light mica particles from the heavier gangue particles. Crushing and grinding of the ore is limited to the least amount necessary to liberate the mica from the host rock. The process is effective in treating mica-bearing ores down to approximately 65 mesh.
A generalized flow diagram of the Bureaus pneumatic concentration method for mica recovery is shown in figure 1. For this study, three types of laboratory-size ore crushers were employed to liberate the mica: a standard jaw crusher, a roll crusher, and a hammer mill. Material too large to be fed to the crushers was broken with a sledge hammer to a usable size.
The jaw crusher had jaws measuring 5 inches wide by 8 inches high (fig. 2). Pieces of material up to 5 inches in diameter were crushed in this machine to minus 1-inch diameter for processing in the air classifier. The roll crusher (fig. 3) had two smooth rolls measuring 5 inches wide and 8 inches in diameter. Pieces of material up to 4 inches in diameter were fed to the crusher and crushed to minus 1 inch. The hammer mill (fig. 4) had a 12-inch-diameter rotating drum with free swinging hammers projecting 3 inches from the drum. The maximum feed size was approximately 8 inches.
The Bureaus pneumatic concentration method for mica recovery is designed to process closely sized particles of mica ore. Two screening units and a two-stage zigzag air classifier are used to process each individual size fraction. The oversize particles of the first screen pass through the air classifier to separate the liberated mica from the host rock. A diagram of the two-stage zigzag air classifier is shown in figure 5.
The ore enters the rougher zigzag section through a rotating air lock. Airflow through the classifier can be varied, depending upon the size of the particles being separated. The airflows used in this study ranged fromapproximately 30 cubic feet per minute for minus35- plus 65-mesh material to approximately 160 cubic feet per minute for plus 4-mesh material. The heavier
gangue material falls through the airstream of the rougher zigzag section to be discarded as tailings. The light mica flakes are carried by the air-stream and are collected in the cyclone on the right side of the figure. This rougher concentrate is fed to the cleaner zigzag section through another rotating air lock. Again, the mica particles are carried by the airstream and are collected in the left cyclone. The cleaner concentrate leaves the cyclone through a third rotating air lock and is rescreened to remove undersize material that was missed by the first screening. The final product is generally a high-grade mica concentrate. Airflow through the cleaner zigzag section is set slightly lower than the rougher zigzag section to produce a purer product. Most of the gangue particles that are accidently carried in the air-stream of the rougher section fall through the cleaner section and are recycled to the rougher section. The screen undersize products are combined and feed the screens and zigzag classifier of the next smaller size fraction.
Industry has not established a standard method of analysis to determine the mica content of a sample. In this study, three methods were used individually and in combination to determine the mica content of the various products. These methods were (1) hand sorting, especially of coarse materials, (2) the inclined plane or cardboard method, and (3) separation in heavy liquids. Analyses were made by physically separating and weighing of the products. Analytical products were examined by binocular microscope to detect misplaced particles. The plus 10-mesh analyses were essentially 100-percent accurate, but were limited by sampling reproducibility. The precision of the analyses decreased as the particle size decreased. A statistical analysis indicated that the measured mica content of the concentrate had a confidence interval of 95 percent plus or minus 5 percent. The same confidence interval for the measured tailings analysis was plus or minus 2 percent. All of the analyses reported in this study should be understood to be within those boundaries of error.
A sample obtained from an Arizona mica-bearing pegmatite contained approximately 22 percent mica. The sample was split into three fractions to study the relative effects of the three types of crushers. Several mica sheets up to about 1.5 square inches of surface area were found in the sample, but most of the mica grains were smaller than 1 square inch of surface area. The sample was run-of-mine rock up to 12 inches in diameter, about 60 weight- percent of which was plus 8 inches. The large size of the pieces made representative samples difficult to obtain, which accounts for the discrepancy in the mica head analyses of the three samples given in tables 1, 2, and 4. The mica was liberated at 4 mesh. The mica contained in the plus 4-mesh material was locked with quartz, plagioclase, and minor amounts of microcline.
A sample of rock was fed to the jaw crusher and screened at 1 inch. The plus 1-inch material was recycled to the jaw crusher and was again screened. These steps were repeated until all material was minus 1 inch. The sample was then fed to the pneumatic concentration system as illustrated in figure 1. The tailings from the initial plus 4-mesh concentration were recycled to the jaw crusher to liberate additional mica and fed to the pneumatic concentrator. The cycle was repeated until no plus 4-mesh tailings remained. This system resulted in a composite concentrate containing 82.1 percent mica with total recovery of 50 percent of the mica in the sample (table 1). The jaw crusher produced a large number of thick, nondelaminated mica flakes that were not recovered during pneumatic concentration. Also, the jaw crusher produced a large number of flat gangue particles that were recovered in the pneumatic concentrator and lowered the grade of the mica concentrate.
A sample of rock was run through the roll crusher prior to feeding to the same pneumatic system. This system produced a composite concentrate containing 92.2 weight-percent mica with recovery of 46 percent (table 2). The roll crusher failed to effectively delaminate the mica flakes, and much mica was left in the coarse tailings. In addition, during the initial treatment of the minus 1-inch, plus 4-mesh material, a small amount of mica was produced that would not respond to pneumatic concentration due to its laminated nature. Recycling this material to the roll crusher tended to cause the rolls to lock and stop running. Hence, this plus 4-mesh book mica was removed and recovered as a separate product. For calculation purposes, it was included in the minus 1-inch, plus 4-mesh concentrate and amounted to 67.7 weight-percent of this size concentrate (table 3).
A sample was run through a hammer mill prior to separation in the same pneumatic system. The hammer mill overcrushed both the mica and gangue. Therefore, the hammer mill was modified by reducing the number of free-swinging hammers from the original 80 to 10, which were spaced about 3 inches apart. Also, the crushing screen or grate was removed so that a particle would receive a minimum number of impacts before leaving the unit. The use of the hammer mill resulted in a concentrate containing 91.5 percent mica with recovery of 70 percent (table 4). On this basis, the hammer mill was determined the best crusher for subsequent research.
A sample of mica-bearing pegmatite from west Texas was obtained for additional studies. The sample contained approximately 9 percent mica with the other constituents being primarily microcline with some quartz and plagioclase. The sample consisted of pieces 3 to 8 inches in diameter containing mica sheets with up to 1 square inch surfaces although most sheets were smaller. Complete liberation of the mica occurred at 4 mesh. The sample was crushed to minus 4 mesh in the modified hammer mill. The sample was then split into two fractions to determine the effects of the hammer mill modification.
One sample was further crushed in an open hammer mill containing 40 hammers andconcentrated. This system produceda concentrate containing 95.3 percent mica witha recovery of 75 percent (table 5).Approximately 9 percent of the mica was lost in the minus 65-mesh material.
The second sample was further crushed in the hammer mill containing 40 hammers with a 1/8-inch crushing screen added. The crushing screen increased the residence time of the ore in the hammer and produced a finer mesh product. Because mica flakes are flexible, they were not broken as finely as the associated gangue, and no additional loss of mica to the minus 65 mesh occurred. This system produced a concentrate containing 92.8 percent mica with recovery of 78 percent (table 6).
A sample of mica schist from west Texas was obtained as part of a study to determine the performance of different types of ores. The sample contained approximately 28 percent muscovite and biotite micas. The other constituents were primarily quartz and plagioclase with minor amounts of microcline, zircon, and tourmaline. As received, the ore consisted of pieces 3 to 10 inches in diameter. Liberation and delamination of the mica was found to occur at 20 mesh.
The schist was crushed in the hammer mill containing 40 hammers with the 1/8-inch crushing screen and then concentrated. This system produced a concentrate containing 92.6 percent mica with recovery of 78 percent (table 7).
Pneumatic concentration has been shown to be a successful method of concentrating mica. The process uses the ratio of surface area to mass to separate the mica from associated materials. The type of crusher used had a decisive influence on the delamination of the mica and greatly affected the results obtained using pneumatic concentration.
The three types of crushers investigated differed substantially in their effects on the pneumatic concentration of the ores. Figure 6 is a photograph comparing the pneumatic concentration products obtained using the three crushers.
Jaw crushers proved to be inferior, producing a concentrate containing only 82.1 percent mica with recovery of only 50 percent. The jaw crusher also produced a large number of thick, nondelaminated mica flakes that did not respond to pneumatic concentration, resulting in a low recovery. In addition, the jaw crusher produced a large number of flat, platy gangue particles that concentrated with the mica flakes, lowering the grade of the concentrate.
The roll crusher produced a concentrate containing 92.2 percent mica with recovery of 46 percent. The roll crusher was not effective in delaminating the mica, which lowered the recovery. The larger pieces of mica jammed the rolls of the crusher and stopped it. Some of the larger pieces of mica could not be delaminated enough to respond to pneumatic concentration. In a continuous system, this material would build up in the crushing circuit. In this study, the nondelaminated mica was removed as a separate product and considered to be a part of the plus 4-mesh concentrate.
The hammer mill was the most acceptable crusher. The pneumatic concentrate contained 91.5 percent mica with recovery of 70 percent. The increased recovery attained by use of the hammer mill resulted from the delamination of the mica particles. Figure 7 is a graph comparing cumulative recovery and mesh size for the three crushers.
Further testing determined the most suitable hammer arrangement to use with the hammer mill. Preliminary testing showed that use of the hammer mill containing 80 hammers resulted in overcrushing of both the mica and gangue. Modifying the hammer mill by reducing the number of hammers to 40 produced a concentrate containing 95.3 percent mica with recovery of 75 percent. The minus 65-mesh fraction was 13.3 weight-percent and contained approximately 9 percent of the mica in the sample.
Placing a 1/8-inch screen in the hammer mill to increase the residence time of the ore produced a finer mesh product. This method of crushing produced a concentrate containing 92.8 percent mica with recovery of 78 percent. The minus 65-mesh fraction increased to 30.1 percent while the total loss of mica remained approximately 9 percent. Because of the flexible nature of the mica flakes, they were not as finely crushed as the associated gangue. The mica flakes were more delaminated with a resultant increase in recovery.
To determine its applicability, this crushing scheme and pneumatic separation was applied to a schist sample. Although a different rock type, with much finer grained material, the system performed equally well, producing a concentrate containing 92.6 percent mica with recovery of 78 percent.
The Bureaus pneumatic concentration method for recovering mica was effective for coarse mica recovery. Liberated mica as large as 1 inch and down to plus 65 mesh was recovered. Since the minus 65-mesh mica is not recovered, the crushing circuit should be designed to minimize the amount of minus 65-mesh mica. The hammer mill delaminated the thick mica particles, thereby increasing the mica recovery by the zigzag air classifier. The pneumatic concentration method, with the modified hammer mill, performed equally well with the two types of ore (pegmatite and schist) tested. The mica concentrates contained over 90 percent mica, representing recovery of as much as 78 percent of the mica in the samples.
The four major steps in the production of marketable copper are mining, concentrating, smelting, and refining. In a few instances, however, leaching takes the place of concentrating, smelting, and refining. At present, although considerable leaching and direct-smelting ores are produced, the bulk of the copper ore mined is concentrated.
The milling of copper ores as practiced in the larger concentrators has changed to such an extent that comparatively few of the machines in use at the beginning of the period remain in service today. Primary and secondary crushing by machines of the Blake and gyratory types and intermediate and fine crushing by rolls has survived, but in the grinding field the development of pebble-mill grinding, the substitution of balls for pebbles, and the parallel development of drag-type classifiers have all but eliminated Chilean and Huntington mills. In the concentrating field, machines which effected separations on the basis of difference in specific gravity between copper and gangue minerals have been almost completely replaced by flotation equipment. In the Lake Superior district jigs and tables have, of course, been retained, and in a few concentrators which treat sulphide copper ores tables have been retained owing to unusual conditions at the plants or the smelters that treat the mill concentrates.
The flotation process, which was responsible for the almost complete change in equipment, has also undergone marked changes since its introduction in large-capacity concentrators. Flotation, when first introduced between 1913 and 1916, was used primarily to reduce losses of copper in the fine tailings of gravity plants. From an accessory to gravity methods, flotation very rapidly became a major process and finally, from 1923 to 1927, all but eliminated the gravity method in the treatment of low-grade sulfide copper ores.
The rapid development of ball-mill grinding must also be attributed to the adoption of the flotation process, since it was the incentive for developing grinding methods which produced considerable copper minerals too finely divided for successful recovery by existing gravity methods.
As with gold and other ores, details of practice vary because of differences in the ores or on account of economic considerations. Five figures are presented to illustrate in a general way the concentrating methods employed for treating the different types of ores.
Figure 150 is the flow sheet of one unit of the gravity concentration section of the Calumet & Hecla Conglomerate mill. The sand tailing from the mill is treated by ammonia leaching and the slime tailing by flotation. Since metallic copper is malleable, it cannot be broken and pulverized as can the more friable minerals, and after first picking out the larger lumps or nuggets of copper by hand, crushing is done by steam stamps; pebble mills instead of ball mills are employed for grinding because of the abrasive qualities of the gangue.
Figure 151 is the flow sheet of one section of the Cananea Consolidated Copper Co. mill as it was in 1929. This is a simple straight-flotation process that replaced an earlier combined gravity and bulk-
Figure 152 is the flow sheet of the Miami Copper Co. concentrator as it was in 1932; A shows the crushing plant and B the grinding and flotation units. Figure 153 is the flow sheet of the Miami concentrate re-treatment and filter plants. The ores are composed of chalcocite and pyrite with subordinate amounts of oxidized copper minerals disseminated mainly in a quartz-sericite schist.
pyrite with minor amounts of gold and silver. The bulk concentrates are dewatered and, after additional grinding, are again subjected to flotation. The latter operation produces finished copper concentrate, finished pyrite concentrate, and middlings which are re-treated.
In its pure form or as an alloy, copper (Cu) is one of the most important metals in society. The pure metal has a face-centred cubic crystal structure, and there is no critical temperature at which this crystal structure changes. Consequently, it is ductile and possesses a high level of electrical and thermal conductivity, making it attractive for a wide range of ornamental and practical applications. With cold-working, copper becomes harder, but it can be made soft again with the heat treating process known as annealing.
Copper was discovered and first used during the Neolithic Period, or New Stone Age. Though the exact time of this discovery will probably never be known, it is believed to have been about 8000 bce. Copper is found in the free metallic state in nature; this native copper is the material that humans employed as a substitute for stone. From it they fashioned crude hammers and knives and, later, other utensils. The malleability of the material made it relatively simple to shape implements by beating the metal. Pounding hardened the copper so that more durable edges resulted; the bright reddish colour of the metal and its durability made it highly prized.
The search for copper during this early period led to the discovery and working of deposits of native copper. Sometime after 6000 bce the discovery was made that the metal could be melted in the campfire and cast into the desired shape. Then followed the discovery of the relation of metallic copper to copper-bearing rock and the possibility of reducing ores to the metal by the use of fire and charcoal. This was the dawn of the metallic age and the birth of metallurgy.
The early development of copper probably was most advanced in Egypt. As early as 5000 bce, copper weapons and implements were left in graves for the use of the dead. Definite records have been found of the working of copper mines on the Sinai Peninsula about 3800 bce, and the discovery of crucibles at these mines indicates that the art of extracting the metal included some refining. Copper was hammered into thin sheets, and the sheets were formed into pipes and other objects. During this period bronze first appeared. The oldest known piece of this material is a bronze rod found in the pyramid at Maydm (Medum), near Memphis in Egypt, the date of origin being generally accepted as about 3700 bce.
Bronze, an alloy of copper and tin, is both harder and tougher than either; it was widely employed to fashion weapons and objects of art. The period of its extensive and characteristic use has been designated the Bronze Age. From Egypt the use of bronze rapidly spread over the Mediterranean area: to Crete in 3000 bce, to Sicily in 2500 bce, to France and other parts of Europe in 2000 bce, and to Britain and the Scandinavian area in 1800 bce.
About 3000 bce copper was produced extensively on the island of Cyprus. The copper deposits there were highly prized by the successive masters of the islandEgyptians, Assyrians, Phoenicians, Greeks, Persians, and Romans. Cyprus was almost the sole source of copper to the Romans, who called it aes cyprium (ore of Cyprus), which was shortened to cyprium and later corrupted to cuprum, from which comes the English name copper. The first two letters of the Latin name constitute the chemical symbol Cu.
When copper and bronze were first used in Asia is not known. The epics of the Shujing mention the use of copper in China as early as 2500 bce, but nothing is known of the state of the art at that time or of the use of the metal prior to that time. Bronze vessels of great beauty made during the Shang dynasty, 17661122 bce, have been found, indicating an advanced art. The source of the metals, however, is unknown.
The Copper Age in the Americas probably dawned between 100 and 200 ce. Native copper was mined and used extensively and, though some bronze appeared in South America, its use developed slowly until after the arrival of Columbus and other European explorers. Both North and South America passed more or less directly from the Copper Age into the Iron Age.
As man learned to fashion weapons from iron and steel, copper began to assume another role. Being a durable metal and possessed of great beauty, it was used extensively for household utensils and water pipes and for marine uses and other purposes that required resistance to corrosion. The unusual ability of this metal to conduct electric current accounts for its greatest use today.